UNIVERSITY  OF  CALIFORNIA 
LOS  ANGELES 


GIFT  OF 

Ann  V*   Adams 


MANUAL    OF    MINING 


BASED   ON  THE   COURSE   OF   LECTURES  ON   MINING 

DELIVERED    AT   THE    SCHOOL   OF    MINES 

OF   THE   STATE   OF  COLORADO. 


BY 

M.  C.  IHLSENG,   C.E.,  E.M.,   PH.D., 

Formerly  Dean  of  the  School  of  Mines  of  the  Pennsylvania  State  College? 

AND 

EUGENE    B.   WILSON, 

Mining  and  Metallurgical  Engineer* 


FOURTH  REVISED   AND    ENLARGED    EDITION. 
SECOND    THOUSAND. 


NEW  YORK : 

JOHN  WILEY  &  SONS. 

LONDON  :  CHAPMAN  &  HALL,  LIMITED. 

1907 


Copyright,  1892,  1898,  190$, 

BY 

M.  C.  IHLSENG. 


ROBERT  DRUMMOND,  PRINTER,   NBW  YORK. 


J  .''  Try 


14 

•]T  a  5 
t> 


PREFACE  TO  THE  FOURTH  EDITION. 


BASED  originally  upon  lectures  delivered  at  the  Colorado 
State  School  of  Mines,  between  1880  and  1888,  the  various 
editions  received  only  such  additions  and  changes  as  were  neces- 
sary in  its  use  as  a  text-book  in  the  mining  schools  of  this 
and  other  countries.  The  Author  of  the  MANUAL  OF  MINING 
presents  in  this  edition  a  complete  revision  of  the  work,  which 
now  assumes  the  form  of  a  collaboration  with  the  authors  men- 
tioned below. 

Mr.  E.  B.  Wilson,  who  needs  no  introduction  to  the  mining 
public,  brings  to  the  original  work  a  wide  range  of  experience 
and  ability,  an  accuracy  of  detail  and  a  discrimination  in  de- 
scription, contributing  highly  to  the  value  of  the  revision.  He 
has  devoted  himself  particularly  to  Chapter  III  of  Part  I  and 
Chapters  V,  VI,  VII,  and  VIII,  of  Part  II.  Originally  planned 
for  Metal-mining,  the  first  issue  of  the  book  gave  small  place 
to  Coal  and  its  extraction.  This  new  edition  has  of  necessity 
been  enlarged  to  include  Coal-mining  in  all  its  phases,  with 
full  descriptions  and  many  illustrations  of  modern  methods  and 
machinery.  The  later  devices  in  power  generation  and  dis- 
tribution, as  the  steam-turbines,  oil,  and  compressed-air  engines, 
with  such  appliances  as  have  proven  themselves  meritorious 
in  economy  and  safety,  are  elaborately  treated.  The  chap- 


IV  PREFACE   TO   THE  FOURTH  EDITION. 

ters  upon  Electricity  in  its  Application  to  the  Mining  Industry, 
prepared  by  Mr.  Rolin  W.  Hutchinson,  Jr.,  whose  work  on  "  Long- 
distance Transmission  of  Electricity"  has  secured  recognition  in 
this  field,  will  prove  invaluable  to  the  student. 

The  Author  desires  to  acknowledge  the  courtesy  of  the  follow- 
ing parties  for  use  of  illustrations  or  tables  as  below: 

Crocker- Wheeler  Electric  Company,  Fig.  66. 
Cyclone  Manufacturing  Company,  Figs.  340,  341. 
De  La  Valle  Steam  Turbine  Co.,  Figs.  48,  49. 
"The  Engineer,"  Chicago,  Figs.  39,  40,  45,  142,  143. 
General  Electric  Company,  Fig.  149. 

Prof.  F.  R.  Hutton,  Mechanical  Engineering  of  Power  Plants, 
Figs.  38,  41,  47. 

Mr.  H.  C.  Reagan,  Locomotive  Mechanism,  Figs.  42,  46. 
Mr.  Frank  Richards,  Air  Compression,  Tables  pp.  308,  309, 

3H- 

Scientific   Publishing  Company,  New  York  City,  Figs.  185, 
186,  187. 

Scientific  Text-Book  Company,  Figs.  95,  104,  178,  188, 193. 

Westinghouse  Electric  Company,  Figs.  49,  in. 

Worthington  Pump  Company,  Figs.  151,  152. 

MAGNUS  C.  IHLSENG. 

NEW  YORK,  November,  1905. 


PREFACE. 


THIS  treatise  is  an  abbreviation  of  a  course  of  lectures 
upon  mining,  delivered  at  the  School  of  Mines  of  the  State  of 
Colorado,  and  is  issued  with  the  advice  of  its  Board  of  Trustees, 
which  recognizes  the  importance  of  having,  within  a  moderate 
compass,  the  best  information  obtainable  upon  this  subject. 
In  its  presentation  the  writer  has  followed  what  his  own  ex- 
perience has  taught  him  to  be  the  natural  sequence,  and  has 
endeavored  to  introduce  such  matter  as  sixteen  years  of  lectur- 
ing and  field  work  have  suggested  as  requisite.  Part  1  contains 
a  brief  geological  review  and  a  discussion  of  such  points  as  the 
engineer  must  include  in  his  report,  i.e.,  the  preparatory  and 
development  work,  systems  of  mining  and  the  plant  for  power, 
hoisting,  pumping,  and  ventilation.  Part  II  embraces  the 
practice  of  prospecting,  drilling,  blasting,  shafting,  tunnelling, 
and  timbering,  in  addition  to  some  remarks  upon  the  examina- 
tion of  mines. 

The  work  is  designed  as  an  elementary  treatise  for  the  use  of 
those  desiring  a  reference-book.  The  complexity  of  the  subject, 
its  extent,  and  the  variety  of  machines  to  be  described  and  repre- 
sented, demand  an  elaborate  discussion  that  would  fill  several 
quartos.  Descriptions  of  obsolete  and  expensive  systems  or 
machinery  are  relegated  to  the  historical  works  on  mining. 
American  and  foreign  practice  is  described,  and  suggestions  for 
lines  of  future  progress  are  offered  herein.  The  principles  of 
the  construction  and  operation  of  machines  used  in  mining  are 


vi  PREFACE. 

explained  with  a  perspicuity  and  conciseness  compatible  with 
the  field  in  which  this  publication  is  to  be  sown — among  students 
and  mining  men,  to  whom  a  knowledge  of  the  fundamenta  of 
their  work  is  valuable,  but  whose  acquaintance  with  the  theory 
is  slight. 

The  wants  of  the  latter  class  have  been  kept  in  mind,  and 
the  writer  hopes  that  the  manual  may  prove  of  some  benefit 
to  the  intelligent  reader,  of  whom  it  presupposes  an  elementary 
knowledge  of  the  sciences  and  of  the  simple  machines. 

The  author  regrets  his  inability  to  deal  with  the  subject  of 
"electricity  in  mining"  as  it  deserves.  Two  reasons  account 
for  this:  insufficient  data,  as  yet;  and  the  large  space  which  a 
satisfactory  explanation  of  the  principles  would  demand. 

The  writer  would  also  beg  leave  to  say  that  the  literature 
of  mining  and  its  cognate  branches  has  supplied  much  of  the 
material  contained  herein.  References  could  not  be  made  for 
each  hint  obtained,  but  obligations  are  acknowledged  to  the 
authors  of  the  publications  mentioned,  to  which  the  reader  is 
referred  for  further  details.  The  information  has  been  gar- 
nered from  the  best  available  sources  and  condensed.  The 
Engineering  and  Mining  Journal,  the  Colliery  Engineer,  and  the 
transactions  of  the  American  Institute  of  Mining  Engineers  have 
been  copiously  drawn  from,  as  also  the  experience  of  the  prac- 
tical men,  to  a  long  list  of  whom  the  Author  is  indebted  for  many 
courtesies.  Finally,  to  the  manufacturers  and  engineers  thanks 
are  rendered  for  the  use  of  the  electrotypes,  which  have  so  largely 
contributed  to  make  the  work  attractive. 

MAGNUS  C.  IHLSENG. 

GOLDEN,  COLORADO,  Nov.,  1891. 


PREFACE  TO  THE  THIRD  EDITION 


IN  the  present  issue  the  text  has  been  increased  by  more 
than  fifty  pages  in  order  to  introduce  additional  matter  per- 
taining to  the  design  of  cars,  hoisting  appliances,  and  fans,  the 
added  illustrations  having  been  specially  prepared  with  that 
end  in  view.  At  the  end  of  each  chapter  will  be  found  a  list 
of  memoirs  which  have  been  carefully  selected  to  comprehend 
the  latest  literature  on  the  subject.  The  original  intent  to  also 
enter  at  some  length  upon  a  discussion  of  the  application  of 
electricity  to  mining  was  abandoned  after  a  number  of  fruitless 
efforts  to  compress  the  matter  into  the  small  compass  which  it 
should  occupy  in  a  work  of  this  nature. 

It  is  gratifying  to  observe  the  progress  made  in  the  increased 
economy  in  the  utilization  of  power.  The  growth  of  the  young 
giant  among  the  motor  fluids  in  the  anthracite  regions  is  par- 
ticularly noticeable;  while  that  of  electricity  in  its  various  ap- 
plications in  bituminous  mines  is  rapid  and  sure.  It  may  be 
true  that  the  adoption  of  these  motor  agents  by  mining  engineers 
has  been  a  little  slow,  but  this  wise  conservatism  is  dictated  by 
the  knowledge  that  a  single  accident  resulting  from  the  intro- 
duction of  an  innovation  might  precipitate  a  peremptory  legis- 
lative prohibition  which  would  render  the  outlay  entirely  useless, 
and  it  is  with  reluctance  that  they  must  frequently  forego  the 
advantages  of  some  possible  economic  installation. 

It  is  gratifying  to  observe  the  extended  employment  of  arti- 
ficial methods  of  ventilation  in  the  metalliferous  mines  of  the 


viii  PREFACE  TO   THIRD  EDITION. 

West,  and  when  the  American  method  of  "square  sets"  shall 
have  been  supplanted  entirely  by  the  method  of  rock-filling, 
or  of  flushing  with  waste,  the  dangers  to  life  and  property  will 
have  been  further  reduced  and  the  economy  of  mining  materi- 
ally increased. 

Regarding  the  enlargement  of  the  chapter  on  explosives 
and  the  stress  laid  upon  the  necessity  for  an  absolute  prohibi- 
tion of  black  powder  from  the  coal-mines,  I  hope  they  may  be 
fruitful  of  results. 

Finally,  there  has  been  added  at  the  end  of  each  chapter  a 
list  of  the  more  important  memoirs  dealing  with  the  subject- 
matter  of  that  chapter,  and  this,  carried  to  the  close  of  the  year 
1897,  has  brought  the  work  up  to  date.  It  should  prove  of 
good  service  as  a  syllabus. 

The  author  hopes  that  the  book  in  its  revised  form  will  con- 
tribute to  a  better  circulation  of  the  knowledge  of  the  principles 
upon  which  mining  engineering  is  founded. 

MAGNUS  C.  IHLSENG. 

\STATE  COLLEGE,  PA..  Dec.  ao,  1897. 


CONTENTS. 


PART   I. 

MINING  ENGINEERING. 

CHAPTER  I. 

PACK 

PROSPECTING i 

Introductory;  native  metals;  ores;  gangue;  definition  of  a 
vein;  fissure  veins;  mineral  beds;  the  theory  of  vein-filling; 
surface  prospecting;  geological  aids  to  prospecting;  explora- 
tions; divining-rods;  the  proof  of  the  existence  of  a  vein;  the 
extra-lateral  rights;  the  dimensions  of  a  mining  claim;  side 
line  vs.  end  line;  maintaining  possessory  right;  the  application 
for  a  mining  patent;  chaotic  condition  of  the  mining  law. 
References. 

CHAPTER   II. 
PREPARATORY  WORK 18 

Mine  development;  the  buildings  at  the  mine  mouth;  the 
selection  of  a  site ;  choice  of  character  of  entry;  the  preparatory 
work ;  adits ;  cross-cut  tunnels ;  vertical  shafts ;  inclines ;  block- 
ing out  the  ore  body;  dead- work;  lifts  in  metal-mines;  the 
lifts  in  coal-mines;  the  dimensions  of  horizontal  opening;  re- 
serves; pillar  supports  for  mines;  influence  of  cleavage  planes; 
faults;  Schmidt's  rule;  open  cuts;  the  steam-shovel;  hydraulic 
mining;  the  duty  of  giants;  nozzles;  peat;  salt-mines.  Refer- 
ences. 

CHAPTER   III. 
METHODS  OF  MINING 34 

Exploitation;  open-cut  mining;  milling;  stripping;  longwall 
advancing;  longwall  retreating;  comparison  of  systems;  pillar- 


X  CONTENTS. 

PAGE 

and-room  system ;  strength  of  coal ;  influence  of  cleats  upon  the 
resistance  of  coal;  size  of  coal-pillars;  relative  dimensions  of 
pillars  and  rooms;  robbing  pillars;  mining  steep,  thick  coal- 
seams;  systems  for  gaseous  mines;  rock-chu'te  mining;  mining 
the  mammoth  coal-seam;  sqware  work;  flushing;  underhand 
stoping;  overhand  stoping;  square  sets;  top-slicing  and  caving; 
subdrifting  and  caving;  slicing  and  filling.  References. 

CHAPTER   IV. 
POWER  GENERATION, 87 

The  power  plant;  the  boiler  rating;  the  standard  of  boiler 
comparison;  the  feed- water;  boiler  scale;  the  feed- water  heater; 
types  of  boilers;  water-tubular  boilers;  heat  losses;  flue- gases; 
the  air-supply;  calorific  value  of  fuels;  the  rate  of  combustion; 
the  methods  of  firing;  stokers;  fuel  consumption;  weathering 
coals;  liquid  fuels;  draft;  chimneys;  blowers;  boiler  installa- 
tion ;  the  steam-engines ;  classification  of  engines ;  the  behavior 
of  steam;  the  mean  effective  pressure;  table,  value  for  the 
constant,  with  various  cut-offs  for  determining  the  M.E.P.; 
the  use  of  indicators;  indicator  cards;  steam  condensation  in 
the  cylinders;  remedies;  the  compound  engine;  governors; 
automatic  cut-off;  the  sliding  steam- valves ;  the  Corliss  valve; 
the  reversing-link ;  the  condenser;  influence  of  clearance;  the 
speed  of  engines;  the  horse- power  of  a  cylinder;  the  horse- 
power of  an  engine;  table,  the  horse-power  for  each  pound  of 
mean  effective  pressure  per  square  inch  of  piston  area;  the 
steam  consumption  of  an  engine;  table,  the  mean  effective  and 
terminal  pressures  and  steam  consumption  in  Corliss  non-con- 
densing-and  condensing-engines;  setting  the  engine;  the  rotary 
effort;  the  steam-turbine;  the  De  Lavalle  wheel;  the  Parsons 
steam-turbine ;  the  advantages  of  the  steam-turbine ;  oil-engines. 
References. 

CHAPTER  V. 
HOISTING  MACHINERY 127 

Hoisting  machinery;  windlasses;  balancing  the  loads;  whims; 
the  hoisting-engine;  first-  and  second-motion  hoisters;  second- 
motion  engines;  the  brake;  the  unbalanced  load  on  the  engine; 
internal  gear  connection;  friction-engines;  V  friction- wheels; 
friction-clutches;  balancing  the  dead  load;  balancing  by  conical 
drum;  the  conical  drum;  the  fusee;  the  reel;  counterbalance; 
the  Koepe  system  of  winding;  chain  counterbalance;  the 
Camphausen  system;  designing  hoisters;  data  necessary  in 
designing;  formulae  for  determining  the  number  and  duration 


CONTENTS.  xi 

PAGE 

•of  trips;  determining  the  size  of  the  rope;  the  bending  stress  on 
the  rope;  the  available  strength  of  a  rope;  example;  deter- 
mining the  dimensions  of  cylinders ;  the  hoisting  capacity  of  an 
engine;  hoisting  on  inclines;  the  starting  moment  of  an  engine; 
table  of  maximum  and  minimum  rotary  efforts  for  certain 
ratios  of  connecting-rod  to  crank-arm;  determining  the  size  of 
the  drum;  the  general  formulas  for  hoisting  resistances;  table, 
the  starting  moments,  working  loads,  etc.,  with  various  hoisting- 
engines  and  drums;  electric  hoisters;  calculating  the  size  of  wire 
for  a  hoister;  example.  References. 

CHAPTER  VI. 

ELECTRIC  GENERATION  AND  WATER-POWER 166 

The  application  of  electricity  to  mining;  electrical  units; 
electromotive  force;  current  force;  capacity;  electric  energy; 
table,  to  develop  one  mechanical  horse-power  at  the  motor- 
shaft;  Ohm's  law;  circular  mils;  electric  currents;  electric 
generator;  the  winding  of  generators;  theory  of  the  dynamo; 
alternating-current  machines;  the  frequency;  the  double- 
current  generator ;  table  of  properties  of  copper  wire ;  examples ; 
general  wiring  formulae  for  alternating-current  distribution; 
example;  resistances;  the  power  factor;  table  of  values  of  M 
for  wires  18  inches  apart;  electric  systems;  converters;  the 
transmission  difficulties;  the  most  economical  area  of  wire; 
the  allowable  voltage  in  mines;  alternating-current  motors; 
rheostats;  impulse- wheels ;  flow  of  water  through  pipes;  fluid 
friction  in  pipes.  References. 

CHAPTER  VII. 

HOISTING  MACHINERY  AND  UNDERGROUND  CONVEYANCES 196 

Underground  conveyances;  hoist-ropes;  steel- wire  ropes; 
standard  ropes;  the  lay  of  the  rope;  locked-coil  ropes;  flat 
ropes;  taper- ropes;  the  strength  of  a  wire  rope;  table  of 
strength  of  cast-steel  hoisting- ropes  and  their  working  loads; 
the  minimum  radius  of  curvature;  table  of  the  radius  of  curva- 
ture R  or  steel- wire  rope  in  inches  having  i°  deflection  and  i 
pound  tension ;  bending  rope  around  curves  and  influence  upon 
its  durability;  the  working  load;  resistance  due  to  bending; 
elastic  connections;  rope  sockets;  £he  head-frame;  the  height 
of  derricks;  sheave;  cage  indicators ;  signalling;  the  telephone; 
means  of  increasing  the  safety  of  ascent  and  descent ;  detaching- 
hooks;  engine  controllers;  buckets;  kibbles;  shaft-guides; 
skips;  gunboats;  the  slope-carriage;  the  cage;  securing  the  cars 


xii  CONTENTS. 

PAGB 

on  the  cage ;  multiple-deck  cages ;  landing-doors  at  shaft  stations ; 
cage-chairs;  self-dumping  cages;  mine-cars;  car  details;  wheels; 
car  resistance ;  lubricants;  testing  lubricating-oils  by  the  hydrom- 
eter; viscosity  test;  the  flash-point;  fire-proof  oils;  handling 
cars  at  the  tipple;  dumping  cars;  cross-overs;  the  Ramsay  cag- 
ing apparatus;  cradle-dumps;  ladders;  the  man-engine.  Ref- 
erences. 

CHAPTER  VIII. 

UNDERGROUND  HAULAGE  SYSTEMS 257 

The  extent  of  haulage  requirements;  the  grade  and  direction 
of  a  haulage-road;  haulage- wavs  in  metal-mines  and  in  coal- 
mines; track  construction;  the  gauge  of  track;  turnouts; 
curves;  the  degree  of  curvature  of  a  curve;  elevation  of  the 
outer  rail;  formulae;  conveniences  at  landings;  safety  devices 
along  haulage-ways;  catches  at  the  top  of  the  incline;  stops 
along  the  incline;  catches  at  the  bottom  of  the  incline;  the 
choice  of  haulage  systems;  tramming;  animal  haulage;  the  loco- 
motive; the  haulage  capacity  of  a  locomotive;  formulae;  the  trac- 
tive force;  table  of  hauling  capacities  of  steam-locomotives  on 
various  grades;  comparison  of  types  of  locomotives;  the  steam- 
locomotive;  the  compressed-air  locomotive;  table  of  tractive 
efforts  of  compressed-air  locomotive;  the  electric  locomotive; 
details;  the  rheostat;  the  draw-bar  pull  of  an  electric  locomo- 
tive; locomotive  rating;  notation;  storage-battery  locomo- 
tives; examples;  cost  of  electric  haulage;  self-acting  planes; 
engine-planes;  rope-haulage  systems;  the  tail-rope  system; 
the  endless-rope  system;  increasing  the  tension  of  the  drums 
at  the  driving  end;  supporting  the  haulage-rope;  deflecting  the 
main  rope  around  the  curves ;  supporting  the  tail-rope ;  example. 
References. 

CHAPTER   IX. 

WIRE-ROPE  TRANSMISSION 295 

Ropes  for  power  transmission;  rope-driving  systems;  the 
transmission  of  power;  sheaves  and  end  carriages;  maintaining 
uniform  tension;  the  tension  of  the  rope;  the  sag  of  rope; 
formulae;  examples;  aerial  tramways;  the  single-rope  tramway; 
the  ore  carriers;  capacity;  the  cable  supports;  the  Bleichert 
system.  References. 

CHAPTER   X. 

THE  COMPRESSION  OF  AIR 307 

Compressed  air;  table  of  the  weights  of  air  at  various  tempera- 
tures: table  of  absolute  pressures,  boiling-points,  etc.,  at  differ- 


CONTENTS. 


ent  heights  above  the  sea-level ;  free  air;  adiabatic  compression; 
isothermal  compression;  formulae;  table  of  volumes,  mean 
pressures,  temperatures,  etc.,  in  air  compression;  examples; 
air  indicator  cards;  the  work  of  adiabatic  compression;  the 
work  of  isothermal  compression;  table  of  hyperbolic  logarithms; 
table  showing  the  horse-power  required  to  compress  and  deliver 
one  cubic  foot  of  free  air  per  minute  to  various  gauge  pressures ; 
the  mean  resistance  to  compression;  cooling  the  air  during 
compression;  water  cooling;  two-stage  compression;  table  of 
horse-powers  necessary  to  compress  100  cubic  feet  of  free  air 
to  various  pressures,  with  two-,  three-,  and  four- stage  compres- 
sors; diagrams;  the  work  of  compressing  moist  air;  table  of  rise 
in  temperature  and  work  of  compression  of  moist  and  dry  air; 
the  air-compressor;  the  valves  of  the  compressor;  the  effect  of 
the  clearance  space;  the  horse-power  of  the  compressor;  the 
receiver;  transmission  of  compressed  air;  factional  resistance 
in  pipes;  table  of  volumes  and  pressure- head  losses  of  trans- 
mission through  pipes ;  the  loss  of  energy  due  to  friction ;  values 
for  a  and  D*  for  various  diameters  of  pipe;  the  most  economical 
size  of  pipe ;  examples ;  compressed-air  pipes ;  table  of  weights  of 
standard  and  extra-strong  pipes  used  for  compressed-air  haulage- 
plants;  the  power  value  of  compressed  air;  the  work  performed 
while  expanding;  reheating;  the  efficiency  of  compressed  air; 
the  Cummings  system  of  air  transmission.  References. 

CHAPTER  XL 

PUMPING 335 

The  water  seepage  into  mines;  methods  of  unwatering  mines; 
the  hydraulic  ram;  hoisting  water;  mine-pumps;  mine  rod- 
pumps  ;  pump-pipes ;  valves  on  pump-pipes ;  pump- valves ;  the 
working-barrel;  formulas;  the  suction  length;  the  lift-pump; 
lift-pump  plunger;  the  capacity  of  lift-pumps;  force-pump 
rods;  the  Cornish  pump;  the  Cornish  rod;  the  balance-bob; 
the  Cornish  engine;  the  speed  of  the  Cornish  pump;  Cushier's 
double-acting  drive-rod  system;  sinking  pumps;  the  recipro- 
cating pumps;  pump- valves;  single-cylinder  pumps;  the  plunger 
pump ;  the  comparative  merits  of  the  steam-pump ;  the  capacity 
of  a  pump;  examples;  the  duty  of  the  pump;  pressure- regu- 
lators for  pumps;  the  speed  of  the  steam-pump;  the  suction 
height;  motor  fluids  other  than  steam;  the  displacement  air- 
pump;  the  air-lift  pump;  water-pressure  engines;  formulas 
for  pump  calculations;  examples;  power-driven  pumps;  the 
Reidler  pump;  electrically  driven  pumps;  rotary  pumps;  the 
centrifugal  pump;  the  theory  of  the  centrifugal  pump;  the 


XIV  CONTENTS. 

PAG* 

compound  centrifugal  pump;  efficiency  of  the  centrifugal 
pump;  table  of  horse-powers  and  fuel  required  for  two-  and 
three-stage  centrifugal  pumps  for  hydraulic  mining  or  for 
pumping  purposes;  the  pulsometer;  siphons.  References. 

CHAPTER   XII. 
MINE-GASES 386 

Ventilation  as  an  economic  proposition;  comparison  of 
metal-  and  coal-mines;  gases  in  coal-mines;  composition  of  the 
atmosphere;  oxygen;  nitrogen;  carbonic  acid;  sulphuretted 
hydrogen ;  carbonic  oxide ;  light  carburetted  hydrogen  or  marsh 
gas;  the  occlusion  of  gas  in  coal;  outbursts  of  gas  from  the  coal; 
the  effects  of  mine-gases  upon  life  and  flame ;  explosive  gaseous 
mixtures;  black  damp;  after-damp;  treatment  for  asphyxia- 
tion ;  the  force  of  the  explosion ;  the  barometric  relation  of  ex- 
plosions; the  diffusion  of  gases;  testing  mine  air  for  gas;  the 
height  of  Che  flame  in  gas ;  the  height  of  the  cap ;  testing  lamps ; 
the  Shaw  gas-testing  apparatus;  the  amount  of  air  required  for 
combustion;  the  water-gauge;  the  mine  resistance;  the  equiva- 
lent orifice  of  the  mine ;  table  of  the  water-gauge  in  two  mines 
carrying  30,000  and  100,000  cubic  feet  of  air;  table  of  air  per 
minute  for  various  equivalent  mine  orifices.  References. 

CHAPTER   XIII. 

METHODS  OF  VENTILATION 405 

The  ventilation  system;  the  ventilation  of  single  entries; 
ventilation  by  double  entries;  planning  airways;  the  under- 
ground temperature;  natural  ventilation;  the  tension  of  the 
atmosphere;  the  weight  of  air;  formulae;  the  production  of 
draught;  the  motive  column;  the  systems  of  producing  a 
ventilating  current;  furnaces;  the  limits  of  furnace  ventila- 
tions; examples;  types  of  fans;  the  trompe;  pressure-blowers; 
the  Fabry  blower;  the  Baker  rotary  force-fan;  Cooke's  fan;  the 
Lemielle  blower:  the  centrifugal  fans;  open  running- fans;  closed 
running- fans;  the  shutter  regulator;  the  influence  of  shape, 
dimension,  and  speed  of  fan  upon  its  capacity;  automatic  speed- 
and  pressure- recorders ;  comparison  of  fan  and  furnace;  the 
theory  of  the  fan ;  the  equivalent  orifice  of  the  fan ;  the  relation 
between  mine  and  fan;  the  efficiency  of  the  fan;  the  design  of 
a  fan;  table  of  theoretical  water-gauge  depression  in  inches 
for  the  corresponding  peripheral  speeds  in  feet  per  second; 
formulae;  examples.  References. 


CONTENTS.  XV 

CHAPTER   XIV. 

PAGB 

DISTRIBUTION  OF  AIR 44* 

The  effective  motive  column;  the  frictional  resistance  of  air 
in  mines;  shape  and  dimensions  of  airways;  table  of  pressure  in 
pounds  per  square  foot  due  to  the  flow  of  air  through  one  mile  of 
airway ;  table  showing  lengths  of  roads  offering  a  resistance  equal 
to  one  inch  of  water-gauge;  formulae;  calculation  of  the  mine 
resistance;  multiple  air-circuits;  the  process  of  balancing  the 
resistances;  the  power  required  for  ventilation;  examples; 
goaves;  the  velocity  of  the  air-current;  the  anemometer;  the 
efficiency  of  a  ventilating  system ;  the  regulation  of  multiple  cir- 
cuits; structural  appliances  required  for  controlling  the  air- 
current;  the  means  for  deflecting  the  current;  the  stoppings; 
air-crossings,  or  bridges;  overcast;  undercast;  how  should  air- 
bridges be  made  ?  mine-doors ;  regulator-doors ;  the  effect  of  an 
obstruction  in  the  airway;  safety-doors;  brattices;  aid  to 
ventilation;  example.  References. 

CHAPTER   XV. 

THE  ILLUMINATION  OP  MINES 468 

Illumination  in  mines;  illuminants;  the  Davy  safety-lamp; 
defects  of  the  Davy  lamp;  Stephenson's  lamp;  the  Marsaut 
lamp ;  the  Muesler  lamp ;  the  Hepplewite-Gray  lamp ;  the  Dick 
patent  port-hole  lamp;  the  Clifford  lamp;  the  Beard-Mackie 
lamp;  the  Wolf  benzine  safety-lamp;  the  requisites  of  a  safe 
lamp;  bonneted  lamps;  locking  lamps;  lighting  and  relighting 
locked  lamps;  use  of  lamps;  examining  and  testing  lamps; 
cleaning  lamps ;  electric  lights  in  mines ;  examples.  References. 

CHAPTER   XVI. 

ACCIDENTS  IN  MINES 483 

The  inherent  dangers;  comparative  hazard  in  nations;  tables 
of  ratio  of  deaths  to  tonnage  and  number  of  employees ;  the  death 
rate  in  mines;  statutory  provisions  governing  underground 
operations;  table  of  average  number  of  accidents  in  mines  in 
the  United  Kingdom  in  semi-decades;  accidents  due  to  falls; 
remedies;  accidents  due  to  cars,  in  shafts,  etc.;  accidents  from 
use  of  explosives;  gas  explosions;  remedial  measures;  the 
theory  of  coal-dust  explosions;  ankylostomiasis;  underground 
fires;  fires  from  blasting-agents;  spontaneous  combustion; 
extinguishing  fires;  protective  measures ;  rules  for  guidance  after 
explosions;  aerophores.  References. 


xvi  CONTENTS. 

PART   II. 

PRACTICAL  MINING. 
CHAPTER  I. 

PAOB 

SHAFTS 5°3 

Shafts;   how  deep  can  we  mine?   a  list  of  some  of  the  deep 
shafts,  and   mines  of    the  world;    sinking  a  shaft;    extending 
shafts  by  pentice;    the  service  of  shaft  timbers;    the  character 
of  the  timbering  of  shafts;    timbering  a  shaft;    timber  joints; 
shaft  linings;  square-set  timbering;  setting  the  timbers ;  timbei 
ing  in  firm  ground ;  repairing  timbers ;  building  landing-stations, 
forepoling;    wood  tubbing  for  circular  shafts;    masonry  walling 
of  shafts ;  shaft  pillars.     References. 

CHAPTER   II. 

SINKING  IN  RUNNING  GROUND 534 

Excluding  watery  strata;  tubbing;  masonry  curbing;  sink- 
ing through  running  ground;  Triger's  method  by  pneumatic 
tube;  the  Kind  and  Chaudron  process  of  boring  shafts;  the 
trepans;  the  tubbing  and  the  false  bottom;  making  a  water- 
tight joint  at  the  bottom;  Lippman's  drill;  Haase's  system; 
the  freezing  system;  conducting  the  seepage  through  the  walls. 
References. 

CHAPTER   III. 

TIMBERING  ROOMS  AND  GALLERIES 545 

The  service  of  timbering;  the  life  of  timber;  timber  consump- 
tion; elements  of  timbering;  props;  mill-holes;  braced  stulls; 
framed  sets ;  joints ;  the  dimensions  of  sets ;  timbering  roads  in 
large  deposits;  lagging;  the  use  of  iron  under  gound ;  masonry 
lining  for  roads;  arch  centres;  tubular  walling  of  galleries; 
dams ;  timbering  soft  ground ;  the  square  set ;  reinforcing  square 
sets;  objections  to  the  system;  cribs;  protecting  underground 
chambers;  timbering  landings;  timber-cutting  tools;  mine 
timber  framing  machine.  References. 

CHAPTER   IV. 

DRIVING  DRIFTS,  TUNNELS,  AND  GANGWAYS 577 

Levels;  adits  and  levels;  the  alignment  of  a  tunnel;  rock 
characteristics;  driving  in  creviced  rock;  driving  by  air-drills; 
driving  slopes;  driving  tunnels;  the  use  of  supplementary 


CONTENTS.  xvii 

PAGB 

shafts;  driving  a  tunnel;  tunnelling  systems;  comparison  of 
the  four  systems;  tunneling  soft  grounnd;  iron  shields  at  the 
tunnel-face;  the  hydraulic  shield;  the  Anderson  pilot  tube. 
References. 

CHAPTER   V. 

DRILLING-  AND  BORING-MACHINES  FOR  EXPLORATIONS 596 

Classes  of  machines;  prospecting-machines ;  bore-holes; 
Poetsch  method;  percussive  drills;  punch-drill;  oil-well  rig; 
the  jars;  Oennhausen's  chisel;  invention  of  M.  Kind  and 
Chaudron  the  jar  universally  used;  the  drill-tools;  sinker-bar; 
spudding-bit ;  auger-stem;  bull  wheel;  the  Mather  and  Platt 
system;  the  temper  screw;  the  sand-pump ;  recovering  lost  tools; 
the  progress;  hand-boring  of  deep  holes;  casing  a  bore-hole; 
artesian  flows;  moss  box;  the  Keystone  prospecting-machine; 
shooting  the  well;  removing  the  casing;  carbonators;  borts; 
the  diamond-drill;  the  core-barrel;  regulating  the  feed  of  the 
drill;  accidents;  rotary  drills;  the  diamond-bit;  drill-bits; 
removing  the  cuttings  and  the  core ;  the  core ;  deflection  of  the 
drill-hole;  fire-setting  system;  miners'  tools;  rotary  drills; 
hand-boring  machines;  the  Chapman  hydraulic  pipe  rotating 
drill;  the  Davis  calyx  drill.  References. 


CHAPTER  VI. 
MINERS'  TOOLS 615 

The  texture  of  rocks ;  metamorphic  rocks ;  the  massive  rocks ; 
underholing  bituminous  coal ;  shooting  off  the  solid ;  shovels ; 
picks;  j adding;  grip-drill;  grip-drill  illustrated;  the  post-drill; 
portable  power  borers;  anchor-pick  or  poll-pick;  breaking 
dimension  stone;  wedges;  the  plug  and  feather;  hydraulic 
wedges;  gads  or  moils;  blockholing;  the  jumper;  the  drill; 
the  depth  of  holes ;  the  stent ;  removing  the  drillings ;  hammers ; 
single-hand  and  double-hand  work;  itemizing  the  drilling  ex- 
penses; the  blacksmith-shop ;  forge  fuel;  welding;  steel;  harden- 
ing steel;  tempering;  the  correct  temper  for  tools.  References. 

CHAPTER  VII. 

CHANNELERS,  DRILLS,  AND  COAL-CUTTERS 635 

Machine  rock-cutters;  quarrying;  the  quarrying  of  dimension 
stone;  channelers;  the  broach-bit;  lewising;  power  drills;  the 
automatic  steam- valve ;  the  tappet- valve ;  the  little  Rand  giant 
drill;  the  Sergeant  drill;  the  fluid-driven  valve;  eclipse  drill; 
the  Schram  drill;  the  Darlington  drill;  the  drill-rods;  the  drill- 


CONTENTS. 

PAGB 

feed;  cushioning  the  blow;  the  drill  supports;  column  support; 
the  drill  tripod;  progress  and  cost  by  machine-drills;  systems 
of  drilling  holes;  the  centre-cut  system;  American  method  of 
tunneling;  simultaneous  firing ;  drilling  in  benches;  Brain's  radial 
system;  the  continuous  process  of  diamond-drilling;  boring 
headings;  electric  rock-drills;  the  solenoid  drills;  the  box 
electric  drill;  coal-cutting  machines;  the  requirements  of  the 
machine;  the  percussion  machines;  the  breast  machine;  long- 
wall  machines;  shearing  coal-cutters;  the  comparative  advan- 
tages of  coal-cutting  machines;  percussive  machines  versus 
chain  machines;  comparison  of  hand-mining  with  machine- 
work.  References. 

CHAPTER  VIII. 

BLASTING 671 

The  principles  of  blasting;  blasting-agents;  blasting-agents 
of  the  first  class;  explosives;  igniting  and  detonating  explosives; 
black  powder;  methods  of  charging  the  hole  with  powder; 
tamping  and  ignition  of  the  powder;  the  fuse;  the  cap;  pre- 
cautions in  the  handling  of  black  powder;  lewising;  powder 
consumption;  the  manufacture  of  nitroglycerine;  dynamite; 
the  storage  of  nitroglycerine;  methods  of  charging  with  nitro- 
glycerine; the  secondary  explosion  of  powder;  smokeless 
powders;  the  theory  of  blasting;  the  line  of  least  resistance; 
flameless  explosives;  detonators;  simultaneous  firing;  expand- 
ing bits;  the  electric  fuse.  References. 

GLOSSARY  OF  MINING  TERMS 696 

SIGNALLING 711 

USEFUL  INFORMATION 712 

TABLE  OF  WEIGHTS  OF  VARIOUS  SUBSTANCES 714 

EQUIVALENTS  OF  FRENCH  AND  ENGLISH  MEASURES 714 


MANUAL  OF   MINING, 


I. 

MINING  ENGINEERING. 


CHAPTER  I. 

PROSPECTING. 

Introductory. — The  search  for  the  useful  and  precious  min- 
erals has  been  diligently  prosecuted  since  the  early  days  of  civili- 
zation; their  discovery  and  application  have  made  nations  pow- 
erful exponents  in  the  world's  history.  And  nowhere  is  this 
fact  better  exemplified  than  in  our  own  land,  in  the  wonderful 
opening  and  rapid  settlement  of  the  Western  mining  States. 

No  subject  is  more  entrancing,  no  occupation  more  exhila- 
rating, than  mining,  with  its  wonderful  kaleidoscopic  changes. 
In  early  times  excavations  were  made  and  mines  worked  only 
to  a  small  depth  and  in  easy  rock,  and  that,  too,  only  for  sub- 
stances of  high  intrinsic  value,  notwithstanding  the  myriads 
of  slaves  to  furnish  the  labor.  The  attempts  at  systematic  min- 
ing were  few  and  far  between ;  but  since  the  advent  of  the  steam- 
engine,  mining  has  been  acknowledged  an  important  profession, 
requiring  technical  education.  Competition  with  the  whole 
world,  brought  about  by  the  improved  means  of  communica- 
tion, the  paucity  of  bonanzas  and  their  rapid  exhaustion,  compel 


2  MANUAL  OF  MINING. 

a  skilful  utilization  of  all  the  aids  to  a  cheap  extraction  of  our 
immense  wealth. 

The  accessibility  of  the  mine  and  the  vendibillty  of  its  prod- 
uct are  the  ever-ameliorating  features  in  the  mining  history  of 
rations,  districts,  camps,  and  individuals,  gradually  divesting 
mining  of  its  risks  and  rendering  it  more  and  more  akin  to  manu- 
facturing. Each  new  camp,  untrammelled  by  tradition  to  keep 
it  in  the  rut  of  prejudice,  displays  its  genius  for  organization 
and  absorbs  the  latest  devices,  tried  and  true.  Nevertheless,  it 
must  be  admitted  that  in  each  camp  an  adequate  solution  of  the 
problem  involves  intricate  questions  of  environment.  The 
economy  of  mining  is  a  function  of  many  variables,  as  geological 
stratigraphy,  subterraneous  uncertainties,  wages,  water,  timber, 
transportation,  and  treatment.  The  constants  are  few.  The 
proper  relation  of  these  it  is  our  province  herein  to  discuss. 

Hitherto  a  gambling  spirit  has  frequently  controlled  invest- 
ments in  metal  mines.  Speculative  tendencies,  not  technical 
economies,  have  dominated  some  of  our  operators;  their  heavy 
aggregate  outlay  may  have  proven  unprofitable,  for  the  present, 
because  of  salted  mines,  attractive  prospectuses,  or  incompetent 
management.  It  must  be  remembered,  however,  that  they  have 
contributed  to  the  prosperity  of  the  country,  and  at  some  later 
date  their  abandoned  exploitations  will  be  pursued  to.  profit t 
when  the  potential  investment  of  to-day  will  have  been  resolved 
into  future  kinetic  dividends,  the  cost  of  production  being  con- 
tinually on  the  decrease. 

Native  Metals. — The  occurrence  of  the  useful  minerals  in  the 
s:ate  of  native  purity  is  rare.  Still  less  often  are  they  found 
superficially:  they  must  be  delved  for.  In  the  extraction  of 
this  subterraneous  material,  and  its  delivery  to  the  surface,  con- 
sists the  art  of  mining.  The  legal  definition  of  a  mine  includes 
such  "workings  as  must  be  artificially  lighted." 

Gold  and  platinum  are  found  native  in  the  placer  accumu- 
lations of  ancient  and  modern  river-beds,  which  furnish  fully 
75  per  cent  of  the  total  output  of  these  metals.  Gold  occurs 
in  segregated  veins,  alloyed  with  telluirum,  and  always  asso- 


PROSPECTING.  3 

dated  with  pyrites  and  titaniferous  iron;  also  intercalated 
between  the  sheets  of  slate  or  shale,  or  finely  disseminated  in 
eruptive  rocks.  The  only  extensive  native  copper  deposit  is 
the  remarkable  product  of  the  Lake  Superior  region,  where 
the  irregular  masses  are  mined  out  of  the  amygdaloid  trap  and 
sandstone.  Singular  masses  of  metallic  iron  ore  are  found  in 
several  localities,  but  they  are  curiosities  and  casual,  if  not 
meteoric.  Native  silver  is  rare  and  occurs  in  Peru,  Mexico, 
Norway,  and  in  the  Lake  Superior  copper  mines. 

Ores. — With  these  few  exceptions  the  metals  are  found  in 
chemical  union  with  non-metallic  substances,  more  or  less  com- 
pletely segregated  to  constitute  mineral.  Any  accumulation  of 
mineral  of  good  quality  and  in  sufficient  concentration  to  war- 
rant the  expenditure  of  energy  for  its  extraction  is  an  ore.  Mani- 
festly this  is  a  fickle  term,  since  it  depends  for  its  stability  upon 
the  casual  conditions  of  the  market  as  well  as  upon  the  mineral- 
ogical  features. 

The  most  common  substance  is  iron,  entering  as  it  does 
into  almost  all  rocks  and  veins.  Its  most  frequent,  and  value- 
less, combination  is  with  sulphur.  Magnetic  and  specular 
oxide  and  the  carbonate  constitute  the  entire  supply.  These 
occur  as  irregular  masses  in  the  rocks  of  every  geological  age, 
or  in  veins  mixed  with  other  minerals,  but  are  chiefly  in  the 
metamorphic  crystalline,  Archaean  rocks.  Zinc  is  obtained  from 
calamine,  franklinite,  and  blende,  which  are  quite  extensively 
distributed  in  the  Carboniferous  strata.  With  very  few  excep- 
tions,  galena  is  exclusively  the  ore  of  lead  The  carbonate  and 
the  sulphide,  in  the  lower  Silurian  and  Carboniferous  strata,. 
mostly  occur  in  irregular  shoots  and  pockets,  and  rarely  argen- 
tiferous. In  the  older  metamorphic  rocks  the  galena  is  con- 
fined in  fissure  veins  carrying  silver  and  gold.  The  main  supply 
of  silver  is  from  its  minerals,  more  or  less  intimately  associated 
with  other  ores.  Similarly  with  them,  it  has  a  wide  geological 
distribution,  and  is  also  found  "dry"  in  fissures.  Copper,  as 
chalcopyrite,  bornite,  and  cuprite,  is  disseminated  in  and  along 
slates  and  sandstones,  rarely  above  the  Triassic.  Many  galena 


4  MANUAL  OF  MINING. 

veins  in  the  metamorphic  rocks  change  with  depth  to  copper. 
Mercury  comes  from  cinnabar,  which  is  found  in  true  veins  and 
in  contacts.  It  is  not  commonly  encountered.  Tin  has  a 
characteristic  occurrence  in  but  one  form,  as  an  oxide,  and 
only  in  gash  or  segregated  veins,  or  "stockwerke"  of  the  older 
rocks. 

Tin  lodes  are  of  the  segregated  type,  and  gold  or  silver  bear- 
ing, pyrites  and  cassiterite  being  the  common  minerals. 

Millerite  and  pyrrhotite  are  nickeliferous  and  occur  in  gash 
and  segregated  veins,  rarely  deeper  than  500  feet.  Rich  films 
of  genthite  in  talc  veins  often  constitute  a  commercial  supply. 

Manganese  ores  (standard  contains  44  per  cent  of  the  metal) 
are  generally  associated  with  limonite  and  occur  in  pockets 
usually  embedded  in  clay  as  contacts  or  beds  or  permeating 
slates.  Films  of  manganese  appearing  in  moss-like  forms  on 
the  face  of  rock  give  it  the  name  of  "landscape"  rock. 

Mica  is  generally  in  bedded  veins, .  instances  of  contacts  and 
true  lodes  being  rare.  They  are  simply  and  always  dikes  in 
coarse  granite.  Hitherto  only  large  slabs  were  sought,  but  now 
the  fine,  clean  mica  has  a  ready  sale  for  lubrication  and  other 
purposes. 

Phosphate  rocks  for  fertilizers,  the  practical  value  of  which 
is  determined  by  the  amount  of  phosphoric  acid  contained,  are 
found  as  beds  of  irregular  thickness;  veins  or  lodes  transversely 
to  the  strike  of  the  strata,  or  superficial  deposits.  Apatite 
occurs  concretionary  in  a  clay  matrix  between  limestone  and 
clay.  These  are  more  frequent  in  the  Miocene. 

Many  of  the  metals  are  incidentally  obtained  from  their 
mineral  compounds  while  smelting  for  other  metals  with  which 
they  are  associated. 

Gangue. — The  metalliferous  portion  of  a  lode  comprises  only 
a  small  portion  of  its  contents.  The  argentiferous  galena,  bor- 
nite,  blende,  or  their  oxidized  derivatives  in  grains,  pockets, 
or  streaks,  more  or  less  connected,  are  associated  with  a  "gangue" 
of  clay,  quartz,  fiuor,  calc,  or  heavy  spar.  These  earthy  mate- 
rials sometimes  are  intimately  mixed  with  the  mineral,  and  again 


PROSPECTING.  5 

lie  in  layers  contiguous  with  it,  or  the  different  constituents  may 
even  manifest  a  ribbon-banded  structure.  • 

Definition  of  a  Vein. — The  metalliferous  and  earthy  contents 
of  a  deposit  constitute  a  bed  or  a  vein,  and  may  exist  under  such 
circumstances  as  to  render  it  workable.  The  term  vein  is  in- 
tended to  describe  a  regular  unstratified  deposit  in  a  fissure  that 
traverses  the  country  for  a  considerable  distance,  longitudinally 
and  vertically.  The  Supreme  Court  has  defined  it  as  "any  zone 
or  belt  of  mineralized  rock  lying  within  boundaries  clearly  sepa- 
rating it  from  the  surrounding  rock."  This  demands  a  well- 
defined  crevice  of  ready  identification,  and  two  solid  walls  to 
give  it  individuality.  Its  lead  must  be  metalliferous.  A  vein  is 
the  filling  of  a  pre-existing  fissure.  The  term  has  lost  its 
original  significance,  for  formerly,  the  mineral  system  was  sup- 
posed to  bear  a  resemblance  to  the  human  circulatory  system. 
True,  the  fissures  have  originated  during  periods  of  great  dyna- 
mic movement,  producing  folds  and  fissures  which  are  supposed 
to  have  extended  deep  into  the  earth's  crust,  but  the  main  artery 
has  yet  to  be  located.  Though  argentiferous  lead  veins  are 
quite  persistent,  no  evidence  exists  for  the  dogma,  so  tenaciously 
held,  that  they  increase  in  richness  with  depth.  They  may  or 
may  not  become  richer,  or  change,  in  constituents.  Examples 
can  be  cited  for  either  side  of  the  argument.  In  folded  strata 
the  deposit  inclines  to  be  thicker  at  the  ridges,  or  troughs,  and 
thinner  at  the  sides  of  the  folds.  But  this  is  not  generally  the 
case  in  massive  rocks. 

Usually  the  vein  matter  is  crystalline.  It  is  commonly  sepa- 
rated on  either  or  both  walls  from  the  surrounding  rock  by  a 
sheet  of  clay  (called  "selvage"  or  "gouge"),  or  by  other  quite 
distinct  lines  of  demarcation.  The  surface  of  contact  of  the 
deposit  with  the  adjacent  rock  is  called  a  wall,  roof,  or  floor, 
according  to  its  relative  position  to  the  miner.  Not  infrequently 
the  walls  are  polished  surfaces  ("slickensides"),  due  to  grinding 
caused  by  the  slips  during  nature's  contortions.  Sometimes 
portions  of  the  vein  have  slid  on  one  another,  causing  "false 
walls";  therefore  the  miner  is  advised  to  occasionally  break 


6  MANUAL  OF  MINING. 

into  the  walls  to  assure  himself  as  to  the  fact.  On  the  other 
hand,  a  vein  may  have  only  one  or  even  no  wall.  In  the  process 
of  mineralization,  the  original  face  or  faces  of  the  fissure  may 
have  become  disintegrated,  and  all  evidences  of  the  looked-for 
wall  obliterated.  In  such  cases,  economic,  not  geologic,  or  legal 
conditions  define  the  vein. 

Fissure  Veins  belong  to  regions  of  metamorphic  action,  and 
are  the  principal  repositories  of  the  precious  metals.  And  it 
is  a  striking  fact  that  they  are  rarely  found  singly,  rather  in  groups 
of  parallel  veins,  often  in  congeries.  Stockwerke  is  a  term  used 
to  describe  a  condition  of  affairs  in  which  the  country  rock  is 
creviced  in  all  directions,  so  that  the  whole  mass  must  be  mined 
out.  Some  are  filled  with  eruptive  matter,  others  with  vein 
matter,  still  others  were  subsequently  closed  without  any  depo- 
sition. The  mineral  components  are  markedly  dissimilar,  and 
indicate  different  sources.  Those  filled  with  the  same  variety 
of  mineral  were  doubtless  produced  by  contemporaneous  forces. 
Those  fissures  which  interrupt  the  continuity  of  the  older  veins 
are  called  cross-courses.  The  manner  in"  which  the  intersec- 
tions occur  determines  their  relative  age.  Their  absolute  age  is 
not  ascertained,  unless  in  stratified  rock.  Drags  are  more  com- 
mon than  is  supposed,  and  should  not  be  confused  with  intersec- 
tions. The  latter  are  usually  richer,  the  former  not  necessarily 
so,  at  the  point  of  juncture.  Many  of  the  older  veins  are  broken 
and  displaced  by  faults.  Not  only  do  veins  "pinch  and  shoot," 
but  the  pay  streak  will  vary  in  thickness,  plunge  from  wall  to 
wall,  or  split  up  into  numerous  feeders  and  ramifications,  and 
even  disappear  in  a  thread. 

Gash  veins  hold  a  subordinate  position  to  fissures,  but 
they  are  of  small  extent,  and  are  usually  confined  to  a  single 
member  of  the  formation  in  which  they  occur.  Their  habitat 
is  unmetamorphosed  sedimentary  rock.  They  have  no  distinct 
walls  or  gouge,  and  are  unreliable. 

Mineral  Beds.— The  important  sources  of  mineral  are  the 
metalliferous  deposits  which  occur  in  the  sedimentary  strata, 
and  are  termed  beds.  While  the  geologists  may  classify  them, 


PROSPECTING.  7 

the  group  is  sufficiently  identified  by  this  term  for  mining  pur- 
poses. It  includes  deposits,  somewhat  irregular  in  dimensions, 
occurring  in  the  transverse  joints  of  the  rocks;  as  cementing 
material  to  the  remnants  of  shattered  or  insoluble  rock;  as  layers 
conformable  with  the  strata;  as  isolated  impregnations  of  grains 
or  bunches  in  porous  rock;  or  as  a  metasomatic  replacement  of 
porous  rock.  They  may  be  found  similar  to  fissures  in  a  certain 
formation,  then  as  a  blanket  contact  parallel  to  the  stratification, 
to  again  plunge  into  a  lower  series  of  rocks  like  a  fissure,  or  branch 
out  into  a  chamber.  They  are  more  easily  mined,  but  are  less 
persistent  in  depth,  than  veins.  Their  mineral  contents  are  very 
compact,  seldom  crystalline,  and  the  gangue  hardly  distinguish- 
able from  the  country  rock.  The  mineral  is  more  or  less  con- 
centrated along  certain  lines  called  "ore-shoots,"  which  probably 
constituted  the  channels  of  communication  with  the  ultimate 
source.  The  same  is  also  true  of  veins. 

The  Theory  of  Vein-filling. — One  requisite  condition  for 
mineral  deposition  is  a  crevice,  a  porous  or  soluble  rock  conduit 
for  the  fluid  from  which  local  action  has  precipitated  the  mineral. 
Open  cavities  were  not  necessarily  pre-existing,  for  a  vesicular 
rock  would  allow  of  an  easy  flow  to  the  magma,  or  it  might  be 
equally  well  secured  by  dissolving  action  on  the  rock  and  a  sub- 
sequent replacement.  This  is  independent  of  its  geologic  posi- 
tion. In  every  age  are  rocks  which  will  satisfy  this  condition. 
Besides  this,  a  long  train  of  circumstances  has  preceded  the 
vein-formation  involving  dynamic  agencies,  heat  and  meta- 
morphism,  and  even  eruptive  action,  as  important  factors.  These 
disturbances  having  been  often  repeated  through  the  different 
ages,  the  older  rocks  were  more  frequently  shaken  up.  Beyond 
this  no  reason  exists  for  the  prejudice  which  favors  certain  geo- 
logical formations  as  ore-bearing. 

The  geognostical  relations  between  veins  and  their  contents 
are  of  importance  to  the  mining  engineer,  but  our  limited  space 
will  not  admit  of  any  discussion  here.  The  various  works  on 
geology  will  supply  the  information  as  to  the  vagaries  mani- 
fested by  ore  occurrences  and  the  numerous  theories  held.  Some 


8  MANUAL  OF  MINING. 

isolated  examples  exist  under  such  circumstances  as  to  suggest 
the  same  origin  for  the  ore  as  for  the  adjoining  rock  formations. 
Many  of  the  beds  and  veins  have  been  impregnated  by  perco- 
lating waters,  perhaps  at  high  pressure  and  temperature,  con- 
temporaneously with  the  country  rock.  Their  metallic  contend 
may  have  been  carried  in  solution  or  they  may  have  been  in  a 
molten  or  a  gaseous  state  when  the  way  for  their  passage  was 
opened.  This  is  a  matter  for  conjecture,  as  is  also  the  ultimate 
source  of  the  mineral.  The  evidences  frequently  point  to  their 
deposition  as  sulphides,  the  oxidized  forms  being  accounted  for 
by  long-continued  action  of  atmospheric  agencies.  In  the 
presence  of  coal  and  bitumen  in  many  lead  and  zinc  beds  is 
suggested  a  theory  of  cause.  The  "water-line"  theory  has 
served  its  day  and  is  no  longer  tenable.*  The  current  theories 
have  offered  more  or  less  satisfactory  explanation  of  the  genesis 
of  some  of  our  ore  deposits.  Some  one  theory  may  explain 
some  of  the  capricious  examples  of  lodes  or  their  anomalous 
fillings.  But  when  we  find  contiguous  depositions  contrasting 
widely  in  point  of  density;  narrower  parts  of  fissures  filled  by 
denser  or  richer  ores;  superior  minerals  higher  up  than  the  more 
volatile  or  lighter  ones,  even  alternating  with  them,  it  is  impos- 
sible to  advance  a  theory  that  is  specifically  applicable  to  all 
ore  occurrences.  Aqueous  currents  have  apparently  conveyed 
the  minerals  from  their  source  and  by  evaporation  chemical 
decomposition  or  electrolytic  action  deposited  them  simultaneously 
with  or  subsequent  to  the  formation  of  the  cavity  in  which  they 
are  found  The  veins  we  find,  but  not  always  the  silver;  and 
this  inability  to  formulate  a  general  law  by  which  to  locate  the 
hidden  bonanzas  has  led  to  the  compounding  of  the  numerous 
witcheries,  and  divining-rods  of  every  conceivable  form,  for 
imposing  upon  the  credulity  of  the  prospector  who  seeks  a  quicker 
means  of  acquirement  than  is  afforded  by  the  use  of  the  pick, 
shovel,  and  patience. 

There  is  no  particular  angle  of  dip  or  bearing  of  trend  that 
is  universally  favorable  to  rich  veins.  Rules  based  upon  such 
observations  are  local  only.  The  same  may  be  said  as  to  the 


PROSPECTING.  9 

supposed  "live "-ness  of  certain  rocks  to  mineral.  Attempts  to 
formulate  indications  of  "quickening"  mineral  by  associations 
with  general  gangue  matter  or  minerals  have  failed  of  general- 
ization. The  mineral  is  where  you  find  it.  The  Cornishman's 
adage,  "riding  a  zinc  horse  to  fortune,"  has  no  verity  in  this 
country.  Each  locality  has  its  own  peculiarities  of  mineraliza- 
tion, which  the  careful  and  systematic  engineer  will  observe 
and  regard. 

With  the  two  classes  of  rocks,  stratified  and  massive,  are 
coexistent  the  two  classes  of  mineral  deposits,  beds  and  veins. 
Though  many  occurrences  are  of  a  nature  that  admits  of  question 
as  to  classification,  for  mining  purposes  a  sharp  line  of  distinc- 
tion is  not  sought.  Legal  technicalities  have  so  confused  the 
definitions  of  deposits  and  veins  as  to  obliterate  all  semblance 
to  the  original  intent  of  geologists  and  mining  men.  Of  this, 
more  later.  At  present  we  shall  consider  some  rules  to  assist 
the  prospector  in  his  search  for  mineral.  And  while  it  must 
be  admitted  that  many  a  find  has  been  made  through  accident, 
the  existence  of  the  ore  would  be  found  not  to  be  at  variance 
with  the  cumulative  rules  of  geologic  science. 

Accordingly,  the  prospector  will  depend  upon  geological 
data.  In  regions  of  stratified  rock  the-  matter  is  simple.  Coal 
is  found  in  three  geological  horizons,  and  the  presence  or  absence 
of  the  rocks  belonging  thereto  is  indicative  of  the  pros- 
pects. 

The  metals  and  their  minerals  are  distributed,  geologically 
and  geographically,  over  a  large  extent.  The  zinc  ores  in  this 
country  occur  in  the  Carboniferous  and  along  the  Mississippi  valley. 
The  Archaean  and  Silurian  are  most  prolific  of  the  other  ores. 
The  precious  metals  are  chiefly  found  in  the  mountainous  dis- 
tricts, because  the  phenomena  attendant  upon  their  formation 
were  conducive  to  the  filling  of  veins,  and  the  forces  which  gave 
character  to  the  mountain  also  impressed  themselves  upon  the 
vein,  which  is  exposed  to  view  and  subject  to  location.  With- 
out some  such  providential  occurrences  to  change  the  monoto- 
nous topography  of  the  preadamic  surface,  bedded  veins  of  the 


10  MANUAL  OF  MINING. 

stratified   districts   would  have  been   revealed   only  by  boring,. 
while  those  in  massive  rocks  might  never  have  been  formed. 

Surface  prospecting  is  confined,  therefore,  to  the  seeking  for 
an  outcrop.  In  igneous  rock  the  outcrop  is  easily  found.  For, 
unless  the  hill  is  covered  with  slide  rock,  it  is  indicated  by  a 
jutting  ledge  (if  the  vein  matter  is  harder  than  the  country  rock), 
or  by  a  sag  (if  it  is  decomposable).  In  heavy  timber  this  may 
go  unnoticed.  At  high  altitudes  snow  in  the  sags  calls  atten- 
tion to  the  leads. 

The  same  is  true  of  coal,  which  is  located  by  the  terraces 
which  mark  the  outcrop.  The  trend  of  the  terrace,  relative  to 
the  topography  of  the  hill,  gives  a  good  idea  of  the  slope  of  the 
coal.  The  bench  itself  may  give  the  desired  information,  but 
usually  it  will  be  found  that  the  coal  dips  with  the  hill,  when 
the  terrace  or  depression  deflects  outward  toward  the  bottom  of 
the  hill,  and  the  reverse  for  a  coal  dipping  inward,  when  the 
outcrop  will  be  concaved  toward  its  top. 

Substances  foreign  to  the  rock  deserve  notice.  Alternations 
in  the  color  of  the  slide  rock  covering  the  hill  are  good  indications 
of  the  presence  of  oxidizable  minerals  above.  So,  too,  vegeta- 
tion is  a  guide.  Iron  springs  often  accompany  the  outcrop  of 
coal;  the  ochreous  covering  of  the  rocks  and  soil  is  noticeable 
near  some  of  the  anthracite  seams,  and  is  common  in  the  semi- 
bituminous  districts.  Masses  of  highly  oxidized  matter,  broken 
from  the  veins,  compose  what  are  called  "blow-outs,"  and  are 
common  in  galena  regions. 

If  no  evidences  of  outcrop  are  thus  found,  "booming"  may 
disclose  it.  During  winter  or  a  wet  season,  snow  or  water  is 
collected  in  a  reservoir  upon  the  hill,  and,  at  a  convenient  time, 
turned  loose  to  plough  its  way  over  the  soil  in  its  fall.  Many 
a  vein  has  been  thus  discovered  without  great  expense. 

In  stratified  regions  the  order  of  the  geological  series  may 
be  observed,  and  certain  fossils  furnish  the  guide.  Or,  if  the 
prospector  is  examining  new  ground,  he  has  but  to  look  for  min- 
eral in  the  float  on  the  surface  or  in  creek-bed.  The  appearance 
of  material  derived  from  erosion  is  indicative  of  the  character  of 


PROSPECTING.  1 1 

the  rock  from  regions  higher  up.  Therefore  the  bed  of  the 
stream,  or  the  hill  slope,  is  minutely  examined  for  fragments  of 
ore,  or  blossom,  and  followed  as  long  as  mineral  is  found.  If 
the  float  or  shode  boulders  are  pebbly  or  rounded,  or  in  vegetable 
soil,  they  have  come  from  afar  and  the  lode  is  not  at  hand.  If 
the  shode  is  large  and  angular,  it  has  not  come  very  far,  and  the 
discovery  of  a  point  beyond  which  no  float  or  blossom  is  detected 
is  presumptive  evidence  of  approach  to  the  vein.  The  lode  will 
be  found  above  the  point  of  discovery,  and  the  prospector  will  go 
in  the  direction  of  the  drainage  and  thoroughly  search  the  ground. 

In  high  altitudes  the  oxidation  of  the  minerals  in,  and  the 
electric  manifestations  of,  the  vein  outcrops  have  assisted  the 
prospector  by  the  light  playing  over  them.  This  is  of  continued 
occurrence  in  Colorado  above  timber  line,  and  particularly  in 
regions  of  arsenical  veins.  - 

When  found,  the  vein  should  be  examined,  and  its  value 
confirmed  at  several  points;  most  grievous  disappointments 
have  ensued  from  testing  of  the  lode  at  one  point  only.  If  the 
country  is  stratified,  care  is  taken  to  ascertain  all  the  data  of 
thickness,  etc.  Frequently  the  ore  oxidizes  and  rots  away,  to  be 
crushed  by  the  overlying  strata,  showing  only  in  a  small  streak; 
•or  the  outcrop  may  fold  back,  "tail  out,"  and  give  false  impres- 
sions of  great  thickness. 

Geological  Aids  to  Prospecting. — Maps  are  serviceable  as 
showing  the  important  features,  and  a  systematic  plotting  of  all 
data,  geological  and  otherwise,  is  necessary.  Dr.  H.  M.  Chance, 
in  the  Second  Geological  Survey  of  Pennsylvania,  has  an  admir- 
able discussion  on  the  construction  of  geological  cross-sections, 
to  which  the  reader  is  referred.  Prospecting  for  oil  or  gas  is 
speculative.  Drill-holes  must  be  carried  down  to  the  oil-bearing 
sands,  and  there  is  no  surface  guide  to  determine  the  site  for  a 
well  to  be  bored  within  the  known  limits  of  a  given  field. 

Explorations. — If  the  surface  fails  to  reveal  the  mineral  body 
sought,  and  there  remains  reasonable  expectation  of  finding 
it,  a  tunnel,  a  shaft,  or  boring  may  be  resorted  to.  The  two 
former  are  more  expensive  but  safer  guides  than  that  offered 


12  MANUAL  OF  MINING. 

by  boring.  Shafting  is  slower  and  more  costly  than  tunnelling, 
but  more  quickly  reaches  a  flat  seam  at  a  point  suitable  for  devel- 
opment. The  steep  pitching  vein  is  perhaps  best  reached  by  a 
tunnel,  if  the  depth  of  vein  so  gained  is  great  enough  to  com- 
pensate for  the  length  of  tunnel.  The  choice  between  them 
depends  upon  local  conditions.  Both  are  advisable  for  shallow 
explorations,  while  drilling  may  be  employed  for  deep  work. 
The  latter  is  very  commonly  employed  on  account  of  its  cheap- 
ness. But  even  when  it  has  determined  the  data  previously 
doubtful,  the  shaft  or  tunnel  must  subsequently  be  driven.  So 
drilling  has  its  limitation  of  use.  It  is  rarely  employed  as  a 
seeker  for  mineral,  but  merely  to  give  confirmation  to,  and  assist 
in  a  rational  estimate  of,  the  value  of  the  undertaking.  Many 
properties  owe  their  rehabilitation  to  the  results  of  the  diamond- 
drill  exploitation,  and  none  should  be  abandoned  until  after  a 
careful  surface  examination  has  been  made  and  followed  by 
numerous  bore-holes. 

Either  the  punch  or  the  diamond-drill  method  may  be  used 
for  the  boring.  The  former  is  cheaper,  but  the  pulverulent 
material  brought  up  by  the  sludger  is  unsatisfactory;  it  may 
indicate  the  constituents  of  the  rocks  pierced  at  different  depths, 
but  can  give  little  of  its  physical  character  or  dip.  The  diamond- 
drill  core  yields  a  little  more  information,  but  even  its  indications 
are  hardly  trustworthy.  It  affords  an  opportunity  to  identify  the 
rock,  but  some  of  the  soft  strata  are  worn  away  or  the  core  may 
be  turned  in  its  tube,  so  its  revelations  are  not  much  better  than 
those  of  the  sand  of  the  punch-drill.  At  best,  the  results  obtained 
from  either  cutter  are  not  conclusive,  for  it  may  have  just  missed 
the  mineral,  or  have  struck  a  solitary,  small,  soft  chunk  of  ore, 
which  would  supply  cuttings  to  discolor  the  sands  for  a  long 
distance  and  give  amazing  report.  Very  important  deductions 
cannot  be  based  solely  upon  the  indications  of  the  borings. 
Only  after  numerous  holes  have  been  bored  and  a  thorough  sur- 
face examination  has  been  made  can  a  conclusion  be  reached. 

Good,  hard  common-sense,  observation  and  pluck  win,  and 
they  alone.  There  is  no  mystery  about  the  finding  of  mineral 


PROSPECTING.  13, 

deposits.  Nature  is  bountifully  supplied  with  precious  metals 
and  valuable  minerals,  but  her  secrets  are  hid.  Only  the  cumu- 
lative information  of  geological  experience  gives  any  clue  as 
to  the  habitat.  Neither  witchery  nor  magic  charm  can  disclose 
the  whereabouts  of  an  ore  body  or  deposit. 

Divining-rods. — The  wizard  with  the  hazel  wand,  or  the 
spirit  medium  who  is  controlled  by  some  disembodied  Comanche 
chief,  is  an  impostor.  He  affects  a  versatility  and  occult  power 
that  transcends  combined  scientific  knowledge,  but  to  a  paltry 
amount  of  "filthy  lucre"  he  is  not  averse,  when  he  plays  upon 
the  credulity  of  natures  which  are  duped  into  making  extensive 
explorations  upon  his  purported  previsions.  This  would  be 
ludicrous,  were  it  not  also  painful,  to  see  the  number  of  mis- 
guided men  who  have  squandered  hopes  and  possessions  in  their 
search  for  a  short  cut  to  wealth. 

The  Proof  of  a  Vein. — The  discovery  of  mineral  at  the  sur- 
face must  be  followed  up  by  the  disclosure  of  the  vein.  In 
order  to  secure  possessory  title  to  it,  the  United  States  laws 
require  the  proof  of  the  existence  of  a  lode  or  vein.  Test  pits  or 
shafts  are  sunk  to  reveal  the  mineral  in  place,  having  a  definite 
direction  of  outcrop.  A  single  shaft  is  not  regarded  as  sufficient 
evidence  of  a  vein.  The  mineral  must  be  in  place.  Even  if 
disintegrated  near  the  surface,  it  is  still  a  vein  if  a  crevice  pre- 
vails carrying  mineral  matter  between  rock  of  a  nature  and 
origin  different  from  it.  The  U.  S.  statutes  divide  mineral 
ground  into  veins  and  placers  only,  hence  the  presumption  would 
be  that  any  well-defined  metalliferous  crevice  capable  of  ready 
identification  by  the  miner  is  a  vein,  whether  fissure  or  not, — 
only  it  cannot  be  a  placer. 

The  Extra-lateral  Rights. — The  difference  in  the  grants 
under  the  two  cases,  besides  a  difference  in  acreage,  is  that  the 
mining  of  ore  within  placer  ground  is  confined  to  the  vertical 
planes  through  the  boundaries  (sec.  2329,  U.  S.  Rev.  Statutes), 
while  vein  deposits  may  be  pursued  along  their  dip,  "through- 
out the  entire  depth,"  even  if  they  "so  far  depart  from  the  per- 
pendicular" "as  to  extend  outside  of  the  vertical  side  lines  of 


14  MANUAL  OF  MINING. 

the  claim";  and  the  extent  of  the  miner's  right  is  determined 
only  by  the  vertical  planes  through  the  end  lines,  which  should 
therefore  be  properly  drawn.  Extra-lateral  rights  are,  therefore, 
accorded  to  vein  locations. 

The  Dimensions  of  a  Claim. — The  prospector  announces  his 
discover)7  by  recording  the  fact,  together  with  the  data  of  dis- 
covery and  the  direction  of  the  outcrop  of  the  vein,  upon  a  stake 
placed  at  the  property  and  upon  the  record  books  of  the  county. 
This  done,  he  may  proceed  to  work. 

Locations  1500  feet  in  length  are  permitted  upon  the  public 
domain  to  the  discoverer  of  the  lode.  But  for  access  thereto,  and 
for  convenience  of  working,  the  U.  S.  grants,  as  incident  to  the 
principal  feature,  surface  ground  which,  measured  from  the  middle 
of  the  vein,  shall  not  exceed  300  feet  on  either  side.  Some  States 
have  reduced  this  to  150  feet  on  each  side,  while  in  some  Colorado 
counties  only  25  feet  was,  and  75  is,  the  outside  limit.  The 
claim  must  be  essentially  a  parallelogram.  It  may  be  1500 
feet,  or  less,  in  length,  located  substantially  along  the  middle 
of  the  apex,  across  which  are  drawn  two  parallel  end  lines  and 
side  boundaries,  within  the  limit  prescribed,  parallel  in  pairs 
following  the  contortions  of  the  outcrop.  However  else  the  Act 
may  be  vague,  it  certainly  is  not  upon  the  fact  of  the  parallelism 
of  the  exterior  boundaries.  Excessive  locations  are  valid  as  to 
the  legal  limit  and  void  as  to  the  excess. 

Maintaining  Possessory  Right. — It  is  incumbent  upon  the 
locator  to  define  the  boundaries  of  his  claim,  by  placing  stakes 
at  all  corners  and  intersections,  to  notify  others  that  the  ground 
is  entered  upon  and  being  exploited.  These,  with  the  filing  of 
a  location  certificate  in  the  county,  maintain  possessor}'  right 
from  the  moment  of  posting  a  location  notice  of  discover)'  upon 
the  lode.  Within  a  reasonable  time  thereafter,  sixty  days  usually, 
the  locator  is  required  to  sink  a  "discovery"  shaft  at  least  10  feet 
into  the  vein.  This  satisfies  the  regulations  regarding  discovery, 
and  maintains  a  mining  right  against  all  comers  until  the  expira- 
tion of  the  calendar  year. 

From  that  time  on,  an  "assessment"  of  $100  must  be  ex- 


PROSPECTING.  1 5 

pended  annually  as  evidence  of  mining  intent.  A  failure  to 
expend  such  sum  constitutes  a  forfeiture,  by  which  the  claim 
reverts  to  the  public  domain,  and  is  subject  to  relocation. 

A  prospector  is  not  confined  to  a  single  entry  upon  a  dis- 
covered lode.  He  may  appropriate  as  many  claims  as  he  chooses, 
contiguous  or  otherwise,  with  that  of  the  first  discovery.  Upon 
each  1500  feet,  or  less,  of  length  he  must  show  the  intent  to.  mine, 
by  a  discovery  shaft  and  the  assessment  work. 

For  the  development  of  the  mine,  the  annual  assessment 
work  may  be  done  upon  the  surface  or  upon  the  vein,  and  all 
efforts  outside  of  the  limits  of  the  location  with  a  bona-fide  intent 
to  work  the  claim  are  justly  considered  as  if  upon  the  claim — 
as,  for  instance,  development  by  tunnel  instead  of  shaft. 

This  concession  is  further  extended  by  the  U.  S.  Supreme 
Court;  for  where  one  person  owns  several  contiguous  claims 
capable  of  being  advantageously  worked  together,  one  general 
system  of  development  may  be  adopted,  after  the  discovery 
shafts  are  driven.  This  encourages  more  economic  work  and 
subserves  the  best  interests  of  all  concerned. 

The  Application  for  Patent.  —  When  development  work  to 
the  value  of  $500,  exclusive  of  buildings,  is  completed  within 
five  years,  by  an  annual  assessment  of  $100,  he  may  proceed  to 
obtain  the  "patent"  or  legal  title  from  the  government.  For 
this  purpose  an  approved  survey  of  the  property  claimed  by 
him  is  necessary.  An  announcement  of  his  intention  to  apply 
for  patent  is  placed  on  the  property,  in  the  local  land  office  and 
in  the  newspapers,  for  sixty  days.  If  no  opposition  is  filed 
against  this  request  to  the  government,  a  patent  will  be  issued 
to  him  when  he  shall  have  proven  the  previous  facts,  and  that 
he  is  a  citizen  of  the  United  States  and  has  paid  $5  per  acre  for 
the  land  enclosing  the  vein. 

According  to  the  laws,  the  applicant  may  obtain  as  many 
claims  as  he  desires,  provided  he  satisfies  the  requirements  for  the 
annual  assessment  prior  to  making  application  to  the  government. 

Chaotic  Condition  of  the  Law. — The  government  has  been 
exceedingly  liberal  to  its  citizens  in  throwing  open  the  mineral 


1 6  MANUAL  OF  MINING.' 

lands  for  exploration  and  occupation.  The  original  intent  was 
to  legalize  the  possessory  system,  which  had  grown  up  in  the 
absence  of  Federal  legislation,  by  which  the  possessor  claimed 
a  vein  for  mining  purposes  and  was  allowed  the  use  for  right 
of  way  to  the  surface  overlying  his  mine.  This  entailed  the 
granting  of  the  mineral  vein  free  but  the  enclosing  non-mineral 
land  by  sale. 

An  exclusive  right  of  enjoyment  "of  all  veins"  cropping 
inside  of  the  boundaries  is  given  with  the  claim.  If  they  are 
discovered  and  entered  upon  in  adjacent  territory,  the  subse- 
quent locator,  according  to  the  laws  of  the  State  of  Colorado,  may 
have  right  of  way  through  the  cross- vein  to  his  ground  on  the 
other  side  of  the  prior  claim,  but  none  of  the  mineral.  In  every 
case  it  is  intended  that  priority  shall  govern.  Sec.  2326  grants 
to  the  senior  locator  the  mineral  at  the  intersection,  and  to  the 
junior  the  right  of  way  through  it. 

By  the  interpretation  of  the  U.  S.  Statutes,  easement  and 
title  were  clearly  intended  to  be  conveyed  for  all  forms  of  metal- 
liferous deposits,  in  the  use  of  the  terms  "veins,  lodes,  or  rock  in 
place."  The  Act  recognizes  any  mineralized  rock  in  place, 
enclosed  in  the  general  mass  of  the  mountain,  as  a  vein. 

The  Statutes  favored  the  miner  and  assumed  to  cover  all 
lodes  whose  indications  were  sufficiently  marked  for  the  miner 
to  continue  explorations  thereon.  A  crevice,  crevice  matter,  a 
fair  wall,  and  mineral  are  the  essential  conditions. 

It  has  been  seen  that  a  lode  claim,  whether  patented  or  not, 
carries  with  it  all  that  is  beneath  the  surface-ground  claimed, 
with  a  servitude  upon  the  adjoining  territory  obtaining  the  right 
of  following  the  dip  of  the  vein,  and  subject  to  a  like  easement 
granted  to  the  locator  on  adjacent  ground  to  pursue  his  vein 
wherever  it  may  go.  This  obtains  until  some  one  can  show  a 
better  right.  The  common  law  as  to  realty  is  modified  when 
applied  to  mining  property. 

It  has,  however,  happened  that  rulings  were  so  made  and 
construed  that  a  party  may  locate  vacant  ground  and  maintain 
ownership  to  the  mineral  covered  by  it,  unless  it  is  shown  that 


PROSPECTING.  17 

the  mineral  body  belongs  to  a  lode  cropping  elsewhere  within  legally 
claimed  ground.  The  proprietor  who  calmly  continued  work 
upon  his  discovery  found  himself  breaking  into  the  subterraneous 
workings  of  others  who  had  stolen  a  march  on  him.  To  secure 
his  right  he  had  to  bring  action  to  eject.  To  vindicate  his  title 
he  had  to  prove  the  lode  to  be  in  place  and  continuous  from  the 
point  of  his  discovery  to,  into,  and  through  the  ground  of  the 
trespasser.  Failing  to  do  which,  his  claim  was  defeated  and  with, 
it  all  incidents  thereto  attached. 

Naturally  the  train  of  reasoning  led  farther  and  farther  away 
from  the  original  intent  of  the  law  to  reward  the  discoverer  of  an 
apex,  until  the  accepted  idea  is  that,  although  the  "defendant's 
location  may  appear  to  you  to  be  along  the  line  of  the  top,  apex, 
or  outcrop  of  the  vein,  it  cannot  prevail  against  a  senior  location 
on  the  dip  of  the  lode." 

To  what  absurdities  the  law  has  led  us,  by  reason  of  the 
vagarious  interpretations,  the  reader  may  learn  by  referring  to 
Dr.  R.  W.  R.  Raymond's  articles  in  the  Transactions  of  the 
American  Institute  of  Mining  Engineers. 

The  remedy  is  to  repeal  the  present  enactment,  or  else  to  so 
prescribe  and  define  the  subjects  of  the  United  States  grant  that 
purchaser  shall  have  a  warranty  title  to  the  entry.  The  side- 
line law  of  Leadville  is  far  preferable  to  the  present  uncertain 
grant  of  extra-lateral  right.  The  risks  of  mining  are  sufficient 
without  adding  this  unnecessary  one.  It  is  singular  that,  what- 
ever amendments  have  been  proposed  in  recent  years,  they  per- 
sist in  retaining  the  essential  feature  of  the  extra-lateral  right 
depending  upon  the  apex.  So  long  as  this  basis  is  retained  they 
will  do  no  good.  There  is  no  reason  for  longer  maintaining  this 
abnormal  and  indefinable  privilege  in  the  grant  of  mineral  land 
made  by  the  U.  S.  Government. 

Justice  W.  E.  Church,  in  a  concluding  and  conclusive  sen- 
tence of  a  decision,  said:  "The  present  laws  are  a  hotbed  of 
litigation  and  a  fruitful  source  of  error."  Judge  Bradley  de- 
clared them  "imperfect,"  and  those  who  have  had  any  experience 
with  them  will  agree. 


CHAPTER  II. 

PREPARATORY    WORK. 

Mine  Development.— Assuming  that  the  question  "Can  it 
pay?"  has  been  answered  affirmatively,  the  next  feature  to  be 
considered  comprises  the  extent  and  character  of  the  surface 
buildings,  the  nature  of  the  preparatory  work,  the  exploitation 
plan,  and  the  machinery  to  be  installed  for  hoisting,  pumping, 
ventilating,  and  for  the  treatment  of  mineral. 

The  Buildings  at  the  Mine  Mouth. — These  comprise  the 
structure  for  the  power  plant  and  such  additional  buildings  as 
are  needed  for  the  preparation  of  the  mine  product.  If  the  latter 
be  pure,  or  rich  enough  to  be  marketable,  or  if  the  coal  is  to  be 
coked,  a  simple  storage  building  is  sufficient.  If  it  is  to  be  sorted, 
crushed,  or  screened  and  shipped,  a  tipple  or  a  breaker  will  be 
required.  If  the  mineral  requires  treatment  and  the  local  water- 
supply  be  deficient,  or  too  acid,  the  washery,  stamp-mill  or  con- 
centrator will  be  located  elsewhere  and  the  mineral  delivered  to 
it  by  tram. 

Collieries  provide  standing  room  for  a  day's  run-of-mine 
cars,  a  storage  track  for  empties  and  a  side  track  for  handling 
the  railway  cars,  wagons,  or  coke-larries.  If  the  coal  is  screened, 
three  or  four  tracks,  side  by  side,  are  connected  with  the  main 
track  at  both  ends. 

All  buildings  should  be  as  nearly  fire-proof  as  the  risk  demands 
and  the  first  cost  will  permit.  A  large  colliery  with,  say,  1000 
acres  of  ground  to  work,  or  a  well-developed  metal-mine,  will 
thus  erect  very  substantial  buildings.  These  should  be  as  near 
to  the  shaft  as  safety  to  the  mine  will  allow  and  yet  be  remote 

18 


PREPARATORY   WORK.  19 

from  the  mouth,  as  far  as  is  consistent  with  economy.  The 
hoisting  engineer  should  have  an  unobstructed  view  of  the  dump- 
ing platform,  and  his  engine  should  be  far  enough  away  to  give 
the  rope  a  good  fleet  angle.  The  hoist-frame  is  open  and  not 
housed. 

The  Selection  of  a  Site.— The  location  for  the  offices,  the 
buildings,  and  the  point  of  shipment  for  the  mineral  must  satisfy 
the  surface  conditions,  must  be  convenient  to  the  shipping- 
tracks,  and  at  the  same  time  one  which  does  not  require  much 
grading  for  the  buildings.  The  location  of  the  railroad  and  its 
switches  fixes  the  location  of  the  ore-sorting  or  coal-screening  build- 
ing. As  the  surface  plant  may  become  quite  extensive,  its  loca- 
tion must  be  considered  with  a  view  to  the  economy  of  handling 
in  such  a  manner  that  the  mineral  shall  descend  naturally  in  its 
progress  through  the  works.  The  location  of  the  screens  fixes 
the  site  for  the  mouth  of  the  mine.  The  latter  must  therefore 
be  in  the  best  position  for  both  the  track  and  the  mine. 

The  opening  into  the  ore  body  must  satisfy  the  underground 
conditions  of  haulage  and  drainage,  and  it  must  be  centrally 
located  to  the  lowest  point  of  the  vein.  It  would  be  an  ideal 
condition  if  the  shaft  which  would  satisfy  the  shipping  require- 
ments could  also  be  placed  in  the  centre  of  the  basin  with  the 
haulage  and  drainage  toward  it  from  all  sides.  As  this  is  not 
likely,  some  compromise  must  be  made  between  the  various 
demands  of  surface  construction  and  mine  operation. 

Choice  of  Character  of  Entry. — The  method  of  attacking  the 
vein,  or  bed,  varies  with  the  inclination  of  the  mineral  body  and 
its  depth.  Undoubtedly,  when  the  mineral  occurs  at  considerable 
depth,  the  sole  method  of  attack  is  by  a  vertical  shaft.  If  the 
depth  be  not  very  great,  a  slope  from  some  convenient  point  may 
be  preferable.  When  the  vein,  or  the  coal-seam,  crops  out  at 
the  surface,  within  the  property  lines,  an  entry  to  the  lower  depths 
is  obtained  by  a  slope  following  the  inclination  of  the  vein.  If 
the  outcrop  so  exposed  is  quite  long  and  extends  along  the  side 
of  the  hill,  several  slopes  may  be  driven  to  advantage.  When 
the  outcrop  occurs  on  the  hillside,  the  vein  occupying  a  position 


20  MANUAL  OF  MIX  IXC. 

somewhat  vertical,  undoubtedly  the  best  method  of  entering  the 
vein  consists  in  driving  one  or  more  horizontal  openings,  or 
adits,  into  the  vein  at  various  heights  according  to  the  system  to 
be  employed  in  the  subsequent  explorations. 

Owing  to  the  difficulties  and  the  expense  of  sinking  shafts 
and  of  hoisting  the  material  through  them  later,  conditions 
sometimes  arise  when  it  may  be  preferable  to  drive  a  tunnel  from 
some  point  as  far  below  the  outcrop  as  possible  or  desirable 
and  toward  the  vein  with  the  shortest  entry.  Such  a  tunnel  is 
called  a  cross-cut  or  a  cross-country  tunnel,  and  in  some  regions 
is  known  as  a  rock  tunnel,  the  significance  of  the  titles  being 
easily  understood. 

Certain  conditions  in  coal-mines  warrant  the  driving  of  rock 
slopes  as  a  compromise  between  the  horizontal  tunnel  and  the 
vertical  shaft.  They  are  employed  where  only  a  shallow  shaft 
would  be  required,  but  where  a  high  speed  of  hoist  is  regarded 
as  desirable. 

Naturally  the  endeavor  is  to  secure  the  most  direct  line  to  the 
point  of  exploitation,  within  the  shortest  time  and  at  a  minimum 
cost,  while  at  the  same  time  constructing  a  permanent  opening 
which  would  serve  for  all  the  work  during  the  life  of  the  mine. 
The  opening  of  the  mine  is  near  to  the  mill-building  where  the 
mineral  is  screened,  weighed,  washed,  or  treated,  preparatory 
to  being  marketed.  The  character  of  the  opening  can  be  decided 
only  after  knowing  the  exact  nature  of  the  underground  condi- 
tions of  the  vein.  Its  extent,  thickness,  depth  and  general  inclina- 
tion must  be  known.  Some  clue  to  this  can  be  obtained  by 
prospecting  the  entire  field  with  the  use  of  bore-holes  and  open 
cuts,  and  geological  maps  are  exceedingly  serviceable,  but  in 
metalliferous  districts  it  is  hardly  possible  to  foresee  the  character 
and  lay  of  the  ore  body.  Without  this  full  knowledge,  it  is  likely 
that  the  opening  made  for  the  mine  during  exploration  would 
later  be  abandoned  as  being  in  an  undesirable  position  or  too 
small  to  meet  the  final  demands  upon  it.  In  coal-mines  the 
eccentricities  of  the  vein  are  not  so  pronounced  as  to  give  the 
same  difficulty  to  the  engineer. 


PREPARATORY  WORK.  21 

Coal  seams  have  less  irregularities  than  do  the  metalliferous 
veins.  Their  contents  are  quite  uniform  in  character,  and  the 
thickness  of  the  seam  varies  to  a  less  degree  than  does  a  vein. 
Both  are  found  dislocated  by  faults  or  shattered  by  movements 
of  the  strata.  The  metalliferous  vein,  however,  not  only  changes 
in  thickness  and  inclination  but  its  mineral  contents  will  vary  in 
percentage  and  even  in  chemical  composition.  Frequently  the 
mineral  at  a  lower  depth  bears  no  resemblance  to  that  at  the 
surface. 

The  character  of  the  mine  opening  depends  upon  whether  the 
ore  body  or  bed  of  coal  is  revealed  at  the  surface  or  not,  and 
whether  or  not  the  outcrop  is  included  within  the  boundaries  of 
the  property.  It  may  occur  on  the  hillside;  its  strike  of  the 
outcrop  may  follow  the  hill,  nearly  horizontal;  or  it  may  lie  in 
the  direction  of  the  slope  somewhat  approaching  the  vertical.  If 
there  is  no  outcrop  the  depth  of  the  ore  body  is  known. 

The  Mine  Opening — The  entry  into  the  mine  may  begin  on 
the  mineral  at  its  outcrop  and,  continuing,  follow  it  wherever  it 
may  go;  or  the  opening  may  begin  at  some  point  on  the  surface 
away  from  the  vein  to  intersect  it  at  a  selected  point. 

The  openings  are  horizontal,  vertical,  or  inclined,  the  choice 
between  them  being  determined  by  the  conditions  named  above. 
The  horizontal  openings  comprise  drifts,  adits,  and  tunnels. 
The  vertical  openings  are  designated  as  shafts.  The  inclined 
openings  are  termed  slopes,  or  inclines.  The  open  cuts  are 
either  quarries  or  strippings. 

Metal-mines  are  operated  usually  from  one  entry  or  opening. 
Only  under  certain  conditions  presenting  dangers  of  caving  and 
other  reasons  requiring  the  hasty  removal  of  the  mineral  is  it 
desirable  for  metalliferous  mines  to  have  many  openings.  Coal- 
mines are  required  by  law  to  maintain  two  parallel  openings  at  a 
certain  stated  minimum  distance  apart.  These  are  driven  simul- 
taneously or  upon  the  coal-seams. 

The  Preparatory  Work. — The  operator  of  a  coal-,  clay-,  or 
lead-mine  is  confined,  in  his  choice  of  site  for  the  shaft,  or 
tunnel,  within  the  property  lines.  Metal  miners,  in  regions 


22  MANUAL  OF  MINIXG. 

other  than  the  national  domain,  are  also  confined  by  the  vertical 
planes  through  their  boundary-lines. 

On  the  national  domain  the  miner  is  not  restricted  to  his  own 
property  in  choosing  the  site  or  the  character  ofen  try  for  his 
mine.  His  property  is  bounded  by  vertical  planes  through  the 
end  lines,  but  not  through  the  sides;  and  he  may  therefore 
operate  his  vein  on  an  inclination  to  any  depth  desirable.  He 
is  not  confined  to  his  surface  boundary  lines  when  selecting  a 
site  for  the  point  of  attack  upon  the  vein.  He  may  sink  a  shaft 
from  outside  the  boundary-lines,  or  drive  a  tunnel  from  the  foot 
of  a  hill  to  his  vein,  provided  only  that  no  other  existing  mining 
rights  are  injured  thereby  or  are  interfered  with.  A  greater 
latitude  of  choice  is  therefore  afforded  the  precious  metal  miner 
than  the  'coal-operator.  Under  such  conditions,  the  tendency  is 
naturally  to  drive  the  horizontal  cross-cut  tunnel  to  the  vein 
as  being  the  most  economical  method  of  developing  the  prop- 
erty. The  great  expense  and  the  changes  in  the  character 
of  the  vein  militate  against  using  a  cross-cut  tunnel,  notwith- 
standing the  facilities  it  offers  for  shipment.  It  is  a  question 
whether  they  would  outweigh  the  importance  of  "  staying  by  the 
mineral." 

Adits. — Veins  cropping  out  on  the  side  of  a  hill  are  often 
exploited  by  driving  adits,  which  are  merely  horizontal  openings 
following  the  vein.  The  cost  of  driving  is  less  than  that  of  a 
slope  or  a  shaft.  The  cost  of  equipping  is  slight,  and  it  serves  all 
purposes  of  permanent  hauling,  drainage,  and  ventilation,  as 
well  as  during  explorations.  These  adits  are  driven  at  distances 
apart  vertically,  depending  upon  the  thickness  of  the  vein  and 
the  facilities  for  shipment.  They  block  the  vein  into  sections, 
each  one  of  which  above  the  adit  can  be  mined  at  a  very  low 
cost.  Their  dimensions,  equipment,  timbering,  and  gradient 
are  the  same  as  those  of  underground  galleries. 

Cross-cut  Tunnels. — These  have  all  the  advantages  of  the 
adit  in  furnishing  a  convenient  haulage  and  drainage-way  at.  a 
point  far  below  the  average  of  the  workings.  They  furnish  a 
cheap,  secure,  and  permanent  entry  if  the  mine  has  been  developed. 


PREPARATORY   WORK.  23 

It  is  a  question,  however,  if  a  long  tunnel  should  be  driven  as 
a  means  of  exploration  before  the  mine  has  been  carried  to  the 
corresponding  level  that  the  value  and  character  of  the  mineral 
may  be  known. 

Vertical  Shafts. — These  constitute  the  most  secure,  but  the 
most  expensive,  method  of  entry  into  the  mine.  But  the  shaft 
presents  many  of  the  objections  which  also  obtain  against  the 
tunnel.  The  shaft  is  the  only  means  of  economically  reaching 
a  coal-bed  which  is  not  shallow  and  whose  outcrop  does  not 
occur  within  the  lines  of  the  property.  For  reaching  veins  this 
is  an  uncertain  method  of  attack,  particularly  if  the  vein  changes 
materially  with  depth  in  its  inclination  and  character  of  mineral. 
The  great  expense  of  sinking  a  shaft  requires  it  to  be  carefully 
located  in  a  position  which  is  certain  to  be  that  of  the  permanent 
outlet  of  the  mine. 

If  a  choice  of  position  be  afforded  the  engineer,  the  foot- 
wall  side  of  the  lode  is  usually  selected  as  being  safer  than  the 
hanging-wall  side.  But  each  successive  lower  cross-cut  is 
longer  than  the  one  above  it,  and  if  the  country  rock  be  hard 
and  the  vein  very  steep,  the  cost  of  these  cross-cut  drifts  would 
soon  become  prohibitory.  For  this  reason  a  majority  of  shafts 
are  sunk  from  the  hanging- wall  side  to  intersect  the  vein  at  a 
moderate  depth. 

The  size  of  the  shaft  is  as  great  as  the  capital  will  allow, 
though  the  tendency  is  to  restrict  its  area.  The  dimensions  of 
the  hoistway  depend  upon  those  of  the  mine  cars,  which  in  turn 
depend  upon  the  character  of  the  underground  haulage  systems 
and  the  dimensions  given  to  the  haulage-ways.  The  width  of 
the  shaft  compartment  is  equal  to  the  length  of  the  car  plus  6 
inches  at  each  end.  This  direction  is  the  long  side  of  the.  com- 
partment and  is  laid  parallel  to  the  mine  tracks  and  also  to  the 
straight  track  to  the  tipple.  The  shaft  may  have  only  two  com- 
partments, both  being  used  for  hoisting;  three  compartments,  in 
which  the  third  is  for  timbering,  piping,  etc.;  or  even  more  com- 
partments, the  last  being  for  ventilation.  The  other  dimen- 
sion of  the  shaft  will  be  determined  when  the  decision  has  been 


24  MANUAL  OF  MINING. 

reached  as  to  the  number  of  cars  to  be  run  upon  each  cage  at  a 
time  and  the  number  of  compartments  required.  The  capacity 
of  the  shaft  is  determined  by  the  conditions  discussed  in  Chapter 
V;  the  dimensions  of  the  shaft,  the  method  of  sinking  and 
timbering,  in  Chapter  I  of  Part  II. 

Inclines. — Inclines  and  slopes  are  in  great  favor  for  many 
reasons.  They  usually  follow  the  vein  and  explore  it,  thereby 
giving  a  return  in  the  mineral  recovered  for  the  cost  of  driving 
The  maintenance  of  slopes  is,  however,  higher  than  the  cost 
in  shafts.  Slopes  are  provided  in  coal  regions  when  the  inclina- 
tion is  over  10°  and  the  depth  to  be  reached  not  over  500  feet. 

When  the  vein  is  constant  in  pitch  and  has  no  irregulari- 
ties, everything  favors  its  driving.  If,  however,  the  vein  has 
occasional  enlargements  or  a  varying  pitch,  its  pursuit  becomes 
awkward.  The  question  then  arises  as  to  whether  it  is  desirable 
to  follow  rigidly  the  irregularities,  to  reduce  them  to  uniform 
pitch  or  to  abandon  the  entry  altogether  for  a  shaft  or  tunnel. 
Only  the  local  conditions  will  settle  that. 

Blocking  Out  the  Ore  Body. — The  preparatory  works  are 
far  from  complete  when  the  mineral  body  has  been  reached. 
A  flat  coal-bed  or  a  slightly  inclined  vein  is  treated,  for  the  pur- 
pose of  mining,  as  a  tabular  mass  of  mineral  which  is  to  be  divided 
into  blocks  by  haulage-ways.  The  latter  are  called  galleries,  or 
gangways,  in  coal-mines,  and  levels  or  drifts  in  metal- mines. 
The  product  from  each  block  is  delivered  down  grade  to  its  gang- 
way, whence  it  is  hauled  to  the  slope-shaft  or  tunnel  to  daylight. 
These  gangways  are  as  numerous  and  driven  as  close  together 
as  the  requirements  for  exploitation  will  demand  or  the  means  of 
the  operator  will  permit.  The  more  numerous  they  are  the  more 
quickly  can  each  lift  above  them  be  exhausted.  On  the  other 
hand,  as  they  cost  more  to  excavate  than  the  corresponding 
volume  of  material  in  the  rooms,  this  character  of  work  must 
be  confined  to  its  lowest  limit.  Between  these  two  conflicting 
requirements,  each  mine  presents  its  own  local  conditions. 

Dead-work. — The  gangways,  or  drifts,  partake  of  the  nature 
of  explorations,  and,  being  more  expensive,  and  in  many  cases 


PREPARATORY   WORK.  25 

only  tentative,  are  termed  dead- work.  By  dead- work  is  understood 
any  bore-hole  into  the  vein,  auxiliary  shaft,  drift,  or  gangway 
and  the  narrow  work  of  connections  between  the  rooms.  Though 
primarily  unproductive,'  dead  work  bears  a  vital  relation  to  the 
economy  of  the  mine.  The  location  of  the  gangway  is  a  matter 
of  great  importance.  If  the  mineral  is  of  varying  grade  of  value 
and  soft,  it  is  laid  along  the  foot-wall.  It  is  in  the  country  rock, 
near  the  vein,  if  the  former  is  softer  than  the  vein  matter.  This 
enables  the  seepage  and  the  drainage  to  be  taken  care  of  and 
reduces  the  risk  of  injury  from  subsidence. 

Lifts  in  Metal-mines. — Metalliferous  veins  are  blocked  by 
driving  levels,  or  drifts,  at  distances  of  from  60  to  100  feet  apart, 
measured  along  the  slope.  These  are  started  on  either  side 
from  the  slope  or  from  each  cross-cut  intersection  with  the  vein 
and  carried  along  the  vein.  The  casual  connections  made 
between  the  adjoining  levels  occur  only  when  the  working  of 
a  lower  lift  has  carried  that  opening  to  the  level  above.  The 
ventilation  requirements  in  metalliferous  mines  are  not  suffi- 
ciently urgent  to  require  the  continuous  circuit  of  air  through 
the  various  openings  in  the  mine.  The  height  of  the  lift  between 
the  levels  and  their  ratio  to  the  thickness  of  the  deposit  depend 
more  upon  the  percentage  of  the  mineral  of  the  vein  contents 
than  upon  the  method  of  mining.  In  soft  ore  the  levels  are 
close  together  and  the  lifts  not  greater  than  60  feet.  In  hard 
rock,  in  thin  deposits  and  in  ore  of  low  grade,  the  distance  between 
ihe  levels  is  greater.  The  maximum,  however,  is  100  feet. 

The  Lifts  in  Coal-mines. — The  coal-bed  is  divided  into  lifts  of 
300  feet  to  600  feet  along  the  slope,  the  greater  length  for  flat 
beds.  Double  entries  are  started  upon  an  inclination  from 
shafts,-  or  slopes,  with  such  grade  as  depends  upon  the  system 
of  haulage.  These  gangways  are  driven  side  by  side  in  pairs 
with  an  untouched  rib  of  coal  of  at  least  20  feet  between  them 
(Fig.  7).  Their  direction  may  also  be  indicated,  if  not  deter- 
mined, by  the  cleavage  planes  of  the .  coal.  One  entry  in  each 
pair  of  gangways  serves  as  the  intake  airway  to  its  lift  of  coal, 
and  the  other  for  the  outgoing  air-current. 


26  MANUAL  OF  MINING. 

In  thick  and  steep  coal-seams  the  haulage-way  is  usually 
built  near  the  floor  to  facilitate  the  landing  of  the  cars  and  the 
parallel  returning  airway  similar  in  area,  above  it  or  nearer  the 
roof.  From  these  gangways  the  rooms  are  driven  upward  on 
the  slope,  at  distances  apart  of  about  40  feet,  to  the  gangway 
above. 

The  Dimensions  of  the  Horizontal  Opening. — The  dimen- 
sions are  governed  by  the  sen-ice  which  the  tunnel,  gangway,  or 
adit  is  to  give.  They  are  not  unnecessarily  large,  for,  in  addition 
to  the  great  cost  of  excavating,  there  is  added  the  additional  cost 
of  supporting  the  increased  opening  and  the  expense  of  main- 
taining it.  If  it  is  to  serve  as  a  ventilating  way  only,  its  area  is  as 
great  as  the  requirements  of  the  ventilation  systems  demand, 
the  height  and  width  depending  on  the  special  conditions.  If  it 
is  to  be  a  haulage- way,  a  single  track  is  provided,  unless  the  num- 
ber of  cars  passing  through  it  for  a  given  output  demand  a  double 
track.  The  latter  would  probably  be  used  only  for  an  endless- 
rope  system  of  haulage.  The  height  of  the  gangway  is  made 
just  sufficient  for  the  passage  of  men  through  it  and  only  as  much 
more  as  is  required  to  furnish  the  area  necessary  for  the  ventila- 
tion current,  provision  having  been  made  for  the  probable  inter- 
ference with  the  current  by  moving  cars.  The  width  must  be 
sufficient  to  provide  ample  clearance  space  beyond  the  cars  for 
the  safety  of  the  men  walking  along  the  way. 

Reserves. — The  number  of  blocks  opened  at  a  time  is  only 
enough  to  maintain  a  regular  output.  Whatever  the  method 
may  be,  it  aims  to  concentrate  the  men  as  much  as  possible,  while 
giving  them  working  places  of  ample  size  to  break  down  the 
mineral  over  as  large  an  area  of  free  face  as  possible,  to  reduce 
the  length  and  cost  of  gangways  to  a  minimum,  leaving  them 
open  only  as  long  as  needed,  and  robbing  the  supports  as 
promptly  as  the  rooms  are  abandoned. 

The  distance  to  which  the  gangways  are  carried  beyond  the 
shaft  or  slope  depends  upon  the  relation  between  the  cost  of 
their  maintenance  and  of  the  haul  to  that  of  sinking  an  entirely 
new  entry  some  distance.  A  mile  is  as  far  as  ordinary  condi- 


PREPARATORY   WORK.  2J 

tions  of  haulage  allow;  beyond  that,  a  new  opening  is  neces- 
sary. But  to  maintain  a  constant  output  for  the  demand  of  the 
market,  the  levels,  or  gangways,  are  kept  in  advance  of  the  imme- 
diate places  of  work.  Their  revelations  furnish  a  guide  for 
future  action.  This  is  particularly  desirable  in  precious-metal 
mines,  the  value  of  whose  mineral  contents  is  so  variable.  These 
advance  blocks,  together  with  those  of  exceptionally  high  grade 
or  of  medium  quality,  previously  developed,  constitute  the  mine 
reserves.  When  the  vein  rock  is  soft  and  requires  considerable 
timbering  for  its  maintenance,  this  character  of  development 
work  is  not  carried  far  beyond  the  work  of  mining. 

Pillar  Supports  for  Mines. — All  permanent  ways  in  horizon- 
tal deposits  are  protected  from  the  crush  of  the  roof  to  secure  a 
safe  outlet.  Shafts  are  surrounded  by  pillars  of  untouched  vein 
matter  or  coal,  whose  dimensions  depend  upon  the  depth  to 
the  bottom  of  the  shaft;  haulage- ways  in  coal  have  massive 
pillars  on  each  side  undisturbed  except  for  the  narrow,  short 
openings  connecting  the  rooms  with  the  haulage- ways;  and  the 
rooms  are  protected  by  pillars  of  undisturbed  coal  on  either 
side.  These  reserves  afford  abundant  support  and  are  undis- 
turbed until  the  working  places  they  guard  are  to  be  abandoned. 
In  veins  or  nearly  vertical  beds  where  the  wall  pressure  on 
either  side  does  not  tend  to  close  the  rooms  or  the  haulage-ways, 
sufficient  security  is  attained  by  horizontal  timbers  between  the 
walls  above  the  haulage- way  and  by  untouched  blocks  of  mineral 
alternating  with  the  rooms  (stopes). 

The  Influence  of  Cleavage  Planes  in  Determining  the  Direc- 
tion of  Roads. — All  strata  are  more  or  less  uniformly  creviced, 
with  horizontal  planes  of  growth  and  vertical  joints,  caused  by 
shrinkage.  Coal  masses  are  similarly  divided  by  one  or  more 
sets  of  planes  producing  rhombohedral  coal.  These  cleavage 
planes  in  coal,  called  "cleats,"  facilitate  the  breaking  of  the 
mineral.  The  direction  of  the  rooms  in  coal-mines  is  fixed  by 
the  direction  of  the  cleat,  the  important  gangways  being  carried 
with  the  cleat.  In  hard-coal  regions,  where  explosives  are  ex- 
tensively used,  the  cleat  does  not  materially  assist  the  breaking 


28  MANUAL  OF  MINING. 

of  the  mineral,  and  as  a  consequence  the  rooms  are  not  dependent 
for  their  direction  upon  the  cleavage  planes.  In  steep  bitumi- 
nous seams  cleat  has  less  importance  than  grade  for  haulage-ways. 
Faults. — A  fault  is  a  break  in  the  continuity  of  a  vein  or 
stratum.  During  some  convulsion  of  nature  the  strata  were 
broken  and  in  the  readjustment  of  position  a  vertical  displace- 
ment occurred.  The  plane  of  fracture  may  be  empty  or  it  ha< 
been  filled  with  extraneous  matter.  When  encountered  during 
the  operation  of  mining  the  fault  plane  or  cross-course  is  pierced 
and  its  strike  and  pitch  noted.  On  the  distant  side,  the  char- 
acter of  the  rock  is  examined.  In  a  stratified  country  it  should 
be  easy  to  identify  the  stratum  and  its  geological  position  relative 
to  that  of  the  ore-bearing  stratum.  Thus  the  engineer  may  be 
guided  up  or  down  according  to  the  direction  of  the  throw.  But 
operations  in  massive  rock  assume  a  more  serious  aspect.  On. 
the  distant  side  of  the  dike  or  cross-course  no  geological  aid  is 
available.  Hence,  in  attempting  to  follow  the  prolongation  of 
the  vein  beyond  the  fracture,  the  engineer  must  rely  upon  his 
knowledge  of  the  local  conditions  or  upon  the  average  of  those 
prevailing  in  similar  ore-bearing  fields. 

Schmidt's  Rule. — It  is  a  matter  of  record  that  80  per  cent 
of  intersected  veins  were  heaved,  apparently,  to  the  right  or  left. 
Those  to  the  right  are  twice  as  many  as  those  to  the  left.  Hen- 
wood  also  discovered  that  the  heaving  to  the  side  of  the  greater 
angle  is  five  times  as  common  as  to  the  smaller  angle.  In  every 
district  may  be  found  a  rule  for  finding  the  other  end  of  the  vein, 
but  it  is  purely  of  local  application  and  unreliable.  To  formu- 
late a  general  rule  out  of  these  numerous  and  apparently  eccen- 
tric displacements  would  seem  well-nigh  impossible;  but  Herr 
Schmidt,  in  1810,  offered  a  solution  to  the  problem,  which, 
though  not  infallible,  is  the  best  extant  and  has  done  valuable 
service.  "  When  the  cross-course  dips  away,  after  going  through 
it,  the  drift  is  run  along  its  far  wall  in  a  direction  opposite  to 
that  in  which  the  vein  pitches.  If  it  dips  toward  the  mouth, 
the  drift  is  carried  along  the  far  wall,  to  the  right  or  to  the  left, 
as  the  veins  dip  to  right  or  left."  The  amount  of  the  displace- 


PREPARATORY   WORK.  29 

ment,  i.e.,  the  distance  to  be  drifted  for  the  continuation  of  the 
vein,  cannot  be  premised.  It  varies  between  very  wide  limits, 
and  is  thousands  of  feet  in  many  localities  Kenwood  averages 
the  throw  of  veins  at  16  feet. 

Open  Cuts. — Open  cuts  are  used  to  explore  the  outcrops  of 
veins  and  shallow  mineral  bodies  They  contemplate  '  limited 
explorations  and  a  smaller  output  and  life  as  compared  with 
that  of  a  mine.  They  are  also  employed  for  stripping  the  sur- 
face coverings  of  flat  veins  which  are  close  to  the  surface.  Quar- 
ries are  extensive  open  cuts  for  the  extraction  of  iron  ore,  slate, 
peat,  building-stone,  and  such  large  bodies  of  soft  minerals  as 
would  be  difficult  to  support  by  timbering  in  the  usual  methods 
of  underground  mining.  In  quarries,  large  masses  are  dis- 
engaged at  a  time  and  the  operations  are  conducted  upon  a  large 
face  at  a  comparatively  small  expense.  The  entire  deposit  can 
be  recovered  to  a  moderate  depth  with  a  high  degree  of  profit. 
The  point  of  attack  for  a  quarry  is  the  lowest  consistent  with 
the  requirements  for  transportation.  It  is  operated  in  benches 
and  cut  by  mechanical  picks  or  channeling-machines  (Fig.  16). 
Hoisting  from  the  pit  is  accomplished  by  derrick  and  bucket 
with  a  form  of  suspended  cable-way.  The  pulsometer  or  a 
centrifugal  pump  is  used  for  drainage.  The  depth  to  which  a 
quarry  can  go  is  limited  to,  perhaps,  200  feet,  beyond  which 
some  more  rational  method  must  be  introduced.  It  involves 
some  form  of  underground  work  which  ultimately  becomes  far 
more  expensive  than  the  mining  which  could  have  been  adopted 
prior  to  the  operations  of  quarrying.  Open  cuts  are  always 
to  be  deprecated  where  a  subsequent  underground  work  is  to 
be  pursued. 

The  objection  to  quarrying  is  the  damage  which  is  done  by 
the  influx  and  accumulation  of  surface-waters.  These  give 
trouble  to  the  miners  and  must  be  removed  as  promptly  as  possi- 
ble to  admit  of  sinking  of  the  quarry  as  readily  as  the  output 
demands.  Again,  the  difficulty  of  propping  the  sides  of  the  cave 
is  no  small  matter,  particularly  if  the  deposit  occurs  between 
pronounced  walls  with  an  overhanging  side.  The  Tilly  Foster 


30  MANUAL  OF  MINING. 

mine  was  deepened  only  by  the  blasting  away  of  200,000  tons 
of  the  overhanging  wall.  This  character  of  quarry  work  is 
dangerous  and  not  economical,  but  is  universally  employed 
for  iron  ores.  In  the  case  of  the  diamond  ground  of  the  Kim- 
berley  mines,  the  quarry  was  replaced  by  a  method  very  closely 
approximating  the  long  wall  for  coal  and  caving  of  iron  ore.  The 
surface-waters  accumulating-  in  the  open  quarry,  which  found 
their  way  into  the  mine,  were  constantly  threatening  danger,  and 
the  operators  finally  dug  tunnels  around  the  quarry  to  catch  and 
divert  the  surface-waters. 

The  Steam-shovel. — In  soft  minerals  the  steam-shovel  with 
its  dredger  derrick  is  employed  for  excavation  purposes.  Its 
introduction  in  the  iron-mines  of  Wisconsin  and  Minnesota  has 
enabled  the  marketing  of  iron  ore  of  a  very  low  grade  at  low  cost. 
The  cost  of  quarrying,  crushing,  and  administration  of  the  iron- 
ore  deposits  in  the  northwestern  portion  of  the  United  States 
and  in  British  Columbia  rarely  exceeds  90  per  cent  per  ton,  and  in 
many  cases  is  very  much  lower.  Some  of  the  copper  deposits 
of  British  Columbia  are  also  of  such  phenomenal  size  that  their 
development  by  the  steam-shovel  has  brought  about  the  intro- 
duction of  machinery  for  hydraulic-,  electrical-,  and  steam-power 
purposes  in  one  case  up  to  6000  horse-power.  Air-machine  drills 
of  3^-inch  size  are  used  as  auxiliaries  where  the  rock  is  suffi- 
ciently hard  to  demand  their  employment  upon  the  immense 
boulders  which  are  interspersed  with  the  softer  rocks.  The 
quantity  of  earth  which  can  be  removed  by  a  steam-shovel  varies, 
of  course,  with  the  supply  of  wagons  which  it  has  and  the  length 
of  their  haul.  But  a  45-ton  Bucyrus  shovel  having  a  bucket 
carrying  1.75  cu.  yds.,  excavated  in  a  day  about  1200  bank  yards 
from  a  cut  whose  depth  was  8  feet  and  whose  average  width 
was  22  feet.  In  this  case  automatic  dump- wagons  were  employed, 
and  with  a  haul  averaging  1500  feet  each  way.  Three  wagons 
were  filled  per  minute. 

Hydraulic  Mining. — This  -is  a  species  of  open  work  in  which 
water  is  the  agent  for  cutting  and  transporting  the  gold-bearing 
earth.  From  nozzles  at  the  lower  end  of  pipe  lines,  leading  from 


PREPARATORY   WORK.  31 

reservoirs  located  at  convenient  points  of  great  elevations,  the 
water  is  discharged  at  a  high  velocity  against  the  bank  of  a  stream 
or  an  old  river-bed  to  undermine  it  and  to  wash  the  material  into 
sluices  conveniently  placed  to  receive  the  flow.  The  sluices  are 
built  on  an  inclination,  with  cleats  on  the  floor  and  of  a  width  and 
depth  sufficient  to  carry  the  volume  of  water  and  its  suspended 
material.  The  later  installations  of  hydraulic  plants  use  centrif- 
ugal pumps  (Fig.  122).  In  two  or  three  steps  the  water  is  raised 
from  the  creek  and  delivered  to  the  bank  at  high  pressure  and 
velocity. 

Hydraulic  mining  is  inexpensive  because  of  the  small  amount 
of  labor  involved.  Four  men  can  operate  the  entire  plant, 
hence  the  process  can  be  conducted  at  low  cost  and  with  a  high 
degree  of  efficiency.  Earth  carrying  6  cents  in  gold  per  cubic 
yard  is  regarded  as  "  pay-dirt."  The  total  cost  of  operation  is 
$0.082  per  miner's  inch.  About  2^  to  3  cu.  yds.  are  moved  per 
miner's  inch  of  water  used. 

A  long  sluice  is  unnecessary.  All  the  coarse  gold,  and  nearly 
70  per  cent  of  the  total  fine  gold,  recovered  is  caught  in  the  first 
section  of  boxes,  about  12  per  cent  being  obtained  in  the  second 
and  an  insignificant  amount  in  the  fifth  section.  The  grade  of 
the  boxes  is  5  inches  per  foot. 

The  Duty  of  Giants  and  Lift-tips. — The  following  table 
shows  the  distances  which  the  standard  nozzles  can  throw  a  stream 
of  water  together  with  the  quantity  of  water  which  they  can  handle 
under  the  stated  pressures. 

Salt-mines. — The  getting  of  salt  is  generally  by  a  special 
process.  It  is  always  found  in  old  river-beds  and  frequently 
quite  thick.  The  process  of  mining  follows  one  of  two  methods. 
Pure  thick  beds  are  mined  systematically  by  a  method  of  squares 
or  like  similar  bodies  of  soft  coal,  due  regard  being  had  to  sound- 
ness of  roof  and  solidity  of  the  mineral.  When  the  salt  body  is 
thin  or  somewhat  impure,  a  different  method  is  employed.  Holes 
are  drilled  to  bed-rock,  water  poured  down  and  the  rock-salt 
bleached  out.  A  pump-pipe  line  is  extended  to  the  floor  level 
and  the  brine  raised  to  the  surface. 


MANUAL  OF  MIXING. 


DUTY  OF  GIANTS. 


Diameter  of 

Nozzle. 
Inches. 

Pressure  at  Nozzle. 

Pressure  in  pounds  
Head  in  feet  

3° 
69-3 

40 
92-4 

5° 
"5-5 

60 
138.6 

70 
IOI-7 

80 
184.8 

90         100 
207.9!    231 

< 

Gallons  .  . 

U4 

9° 
62 

*55 
109 

70 

173 
126 

94 

189 
142 
1  08 

% 

121 

219 
168 
131 

232    1    245 
178     |     186 
140         148 

Horizontal  distance  
Vertical  distance  

I* 

Gallons 

170 
93 
63 

196 

"3 
81 

219 

132 
97 

240 
148 

112 

259 
163 
125 

277 
175 
137 

294 
186 
148 

3i<> 
193 
157 

Horizontal  distance  
Vertical  distance  

ii 

Gallons  
Horizontal  distance  
Vertical  distance  

210 
96 
63 

242 
118 
82 

271 
138 
99 

297 
156 

"5 

320 
172 
I29 

342 
1  86 
142 

9 

154 

383 
207 
164 

i? 

Gallons  
Horizontal  distance  
Vertical  distance  

253 
100 

64 

293 
124 

83 

327 
146 

100 

118 

387 
l84 
133 

413 

200 

I46 

439 
213 
158 

462 
224 
169 

ii 

Gallons 

313 
104 

64 

361 
130 
83 

404 

X52 
101 

442 
176 
119 

478 

*95 
135 

5" 

212 
149 

54i 
224 
163 

571 
240 

175 

Horizontal  distance  
Vertical  distance  

if 

Gallons  

425 

no 

65 

490 
140 
85 

548 

161 
104 

600 
1  88 

121 

649 

2IO 
138 

604 
230 
'54 

735 
243 
169 

775 
259 
183 

Horizontal  distance  
Vertical  distance  

2 

Gallons 

556 

116 

65 

642 

147 
86 

717 
166 

105 

785 
194 
124 

849 

218 
141 

907 
240 
158 

962 
255 
J74 

1014 
274 
189 

Horizontal  distance  
Vertical  distance  

A  nozzle  if  inches  diameter,  under  207.9  ^eet  head,  can  throw 
439  gallons  cf  water  per  minute  158  ft.  high  or  213  feet  distant. 

Peat  is  recovered  by  dredging  channels  usually  cut  in  the 
heavy  bogs  deep  enough  to  serve  for  drainage  and  for  navigation. 
The  peat  is  then  cut  in  blocks.  When  it  occurs  above  water  the 
fuel  is  quarried  in  steps  as  are  building  materials. 


PREPARATORY    WORK.  33 


REFERENCES. 

The  Genesis  of  Certain  Auriferous  Lodes,  J.  R.  Don,  A.  I.  M.  E.,  Vol. 
XXVII,  564;  The  Telluride  Ores  of  Cripple  Creek,  Kalgoorlie,  T.  A.  Rick- 
ard,  A.  I.  M.  E.,  Vol.  XXX,  708;  The  Greatest  Gold-producing  Mines,  J.  H. 
Curie,  Eng.  &  Min.  Jour.,  Nov.  7,  1903;  The  Geologist  in  Matters  of  Prac- 
tical Mining,  J.  E.  Spurr,  Eng.  &  Min.  Jour.,  April  n,  1903;  Mining  Geology, 
G.  J.  Binns,  N.  Z.  Mines  Rec.,  Dec.  16,  1901;  Our  Coal  Resources  and  Con- 
sumption, Profs.  Hall  and  Jevons,  Jour.  Gas  Light.,  Feb.  20,  1900;  The  Dis- 
covery of  New  Gold  Districts,  H.  M.  Chance,  A.  I.  M.  E.,  Vol.  XXIX,  224; 
Connellsville  Coal  Region,  W.  C.  Irvin,  E.  &  M.  Jour.,  Mar.  24,  1900;  The 
Ducktown  Copper  Mining  District,  S.  W.  McCallie,  Eng.  &  Min.  Jour., 
Oct.  4,  1902;  Some  Notes  on  the  Nome  Gold  Region  of  Alaska,  F.  C.  Schrader 
and  A.  H.  Brooks,  A.  I.  M.  E.,  Vol.  XXX,  236;  The  Coal-mining  Industry 
of  the  United  Kingdom,  R.  A.  S.  Redmayne,  Engineering  Mag.,  Dec.  1903. 

The  Prospecting  Density  Rule,  J.  H.  Pollock,  A.  I.  M.  E.,  Vol.  XXIX, 
281;  The  Equipments  of  Camp  Expeditions,  C.  H.  Snow,  A.  I.  M.  E.,  Vol. 
XXIX,  157;  Bibliography  of  Mexican  Geology  and  Mining,  R.  A.  Santillan, 
A.  I.  M.  E.,  Vol.  XXXII,  605;  A  Glossary  of  Spanish  American  Mining  and 
Metallurgical  Terms,  A.  S.  Dwight,  A.  I.  M.  E.,  Vol.  XXXII,  571;  Histori- 
cal Sketch  of  Mining  Legislation  in  Mexico,  E.  M.  Baco,  A.  I.  M.  E.,  Vol. 
XXXII,  520;  A  Synopsis  of  the  Mining  Laws  of  Mexico,  R.  E.  Chism, 
A.  I.  M.  E.,  Vol.  XXXIII,  3;  Mining  Rights,  Foreign:  Germany,  Austria,  Coll. 
Guard.,  Dec.  4,  1896,  1063;  Laws  as  to  Corporations,  etc.,  in  U.  S.  Domain, 
Coll.  Guard.,  July  1897,  105;  Mining  Gold  Ores  in  California,  general  article 
State  Min.  Bureau,  zoth  Report,  1852;  Tunnel  Decisions  in  Colorado,  and 
Editorial,  E.  &  M.  Jour.,  May  22,  1897,  515. 

Our  Coal  Resources  and  Consumption,  Profs.  Edward  Hall  and  Stanley 
Jevons,  Jour.  Gas  Light.,  Feb.  20,  1900;  Development  of  the  Connellsville 
Coal  Region,  W.  G.  Irvin,  Eng.  &  Min.  Jour.,  Mar.  24,  1900. 

Placer  Mining,  Illustrated,  by  A.  Lakes,  Coll.  Guard.,  Eng.,  May  1896,  219; 
Mining  Methods  in  Peru,  S.  of  M.  Quart.,  XV,  219;  Hydraulic  Mining, 
Strength  of  Beams  for  Flumes,  Min.  &  Sci.  Press,  May  29,  1897,  452;  A 
Modern  Dredging  Plant,  Min.  &  Sci.  Press,  Mar.  24,  1900;  Hydraulic 
Sluicing  in  Collingwood  District,  Logan,  N.  Z.  Mines  Rec.,  Feb.  16,  1900; 
and  Hydraulic  Mining,  E.  B.  Wilson. 

Cleavage  Development,  Journal  Geology,  IV,  4  pages,  444. 

Colliery  Surface  Arrangements,  Ernest  H.  Thomas,  Coll.  Guard.,  Mar. 
30,  1900. 

A  Modern  Dredger  Mining  Plant,  Min.   &  Sci.  Press,  Mar.  24,  1900. 


CHAPTER  III. 

METHODS   OF  MINING. 

Exploitation. — The  plan  pursued  for  winning  minerals  and 
making  them  available  for  use  is  termed  exploitation,,  while  the 
general  term  mining  is  applied  to  the  labor  of  excavation. 

For  the  purpose  of  exploitation  mineral  deposits  may  be 
classified  according  to  their  shape  and  inclination  into  those 
which  are  of  irregular  shape  and  those  which  are  tabular  in 
form.  The  former  are  mined  by  open  cut,  quarry,  stripping, 
or  the  hydraulic  process,  with  or  without  the  aid  of  a  dredger  or 
steam-shovel,  as  they  are  usually  superficial  in  position  and 
their  contents  friable  or  soft.  The  latter  are  composed  of  com- 
paratively hard  material  and  are  extended  in  a  horizontal  or 
vertical  direction  with  some  degree  of  uniformity.  Their  con- 
tents are  recovered  by  some  system  of  Underground  Mining. 

The  character  and  depth  of  the  superficial  covering  may 
determine  the  choice  of  method  for  some  deposits  which  occur 
in  tabular  masses.  The  relative  convenience  and  cost  of  remov- 
ing the  cover,  and  the  increased  expense  of  supporting  it  above 
the  subterranean  works,  constitute  the  decisive  element.  Open- 
cut  mining  may  be  profitable  when  the  quantity  of  superficial 
material  to  be  stripped  from  the  surface  does  not  exceed  one 
half  of  a  foot  depth  for  each  foot  of  thickness  of  material  in  the 
deposit,  and  where  the  quality  of  the  cover  is  such  as  can  be  easily 
removed  by  the  steam-shovel. 

Open-cut  Mining. — The  method  of  extracting  the  iron  ore 
in  the  Mesaba  range  in  Minnesota  as  adopted  by  the  Iron  Mining 

34 


METHODS  OF  MINING.  35 

Company  consists  in  stripping  the  surface  material  by  a  steam- 
shovel  of  the  dipper  type.  This  machine  has  a  radius  of  action  of 
15  feet  and  excavates  and  delivers  the  dirt  to  cars  of  the  stand- 
ard railroad  gauge.  A  cut  is  first  made  for  a  certain  distance  and 
width  in  which  the  track  is  laid  to  admit  the  machine.  The 
steam-shovel  excavates  the  dirt  on  one  side  of  the  cut,  delivering 
it  on  the  other  side,  and  stripping  the  surface  for  the  depth  of  cut 
to  the  boundaries  of  the  property.  The  shovel  is  returned  to  its 
starting-point,  and  from  the  side  cut  where  the  dirt  has  been 
stripped  a  new  track  is  laid  and  the  work  proceeds,  while  cars 
are  loaded  on  the  first  track.  The  stripping  advances  over 
the  entire  area  to  the  level  of  the  track.  Several  steam-shovels 
strip  in  a  similar  manner  until  the  ore  body  has  been  exposed. 
The  usual  depth  of  a  cut  is  20  feet;  beyond  that  there  is  danger 
of  the  bank  caving  and  causing  injury  to  men  and  machinery. 
Occasionally  deep  holes  are  drilled  and  chambered  into  the  bank 
ahead  of  the  machine,  into  which  several  kegs  of  powder  are 
loaded  and  blasted  for  the  purpose  of  loosening  the  ore  for  the 
work  of  the  shovel. 

The  attack  upon  the  iron  ore  proceeds  in  the  same  manner 
as  in  the  surface  material — in  benches  of  a  width  of  30  feet,  con- 
taining one  track  for  cars  and  a  short  section  of  track  for  the 
machine.  The  product  per  shift  of  ten  hours  is  1800  tons, 
though  there  are  records  of  a  daily  product  of  3544  tons. 

The  cost  of  stripping  and  wasting  the  mineral  of  a  coal-bed, 
under  the  most  favorable  conditions,  is  approximately  20  cents 
per  cubic  yard,  whether  machines  are  employed  or  hand  labor 
alone  is  utilized  for  extraction.  The  sole  advantage  in  the  intro- 
duction of  machinery  is  the  rapidity  with  which  work  can  be 
prosecuted  and  the  increased  area  of  stripping  which  is  possible. 
These  are  the  determining  features  in  situations  where  the  cover 
is  thick  and  the  working  season  short.  Inasmuch  as  the  rail- 
road approach  to  the  several  benches  on  which  operations  are 
conducted  must  also  be  lowered  with  the  progress  of  the  work, 
it  soons  happen  that  its  grade  becomes  excessive  and  the  traffic 
is  interfered  with  unless  a  corresponding  increase  in  the  length 


36  MANUAL  OF  MINING. 

of  the  approach  be  made.  This  necessitates  a  deep  open  cut 
for  extensive  work,  or  a  tunnel  through  which  the  railroad  cars 
may  be  drawn  to  the  works. 

The  Milling  System. — When  a  limit  has  been  reached  for 
economical  depth  in  the  quarry  an  underground  process  may  be 
introduced  in  combination  with  the  open  cut.  This  is  termed 
milling,  and  has  been  practised  to  advantage  in  the  Lake  Superior 
iron  region.  A  mill  is  a  hole  made  in  the  rock  or  ore  connecting 
an  upper  with  a  lower  level  of  the  deposit,  through  which  material 
is  sent  down  to  the  level  below.  In  the  system  mentioned  above 
the  tunnel  is  driven  large  enough  to  admit  railroad  cars  into  the 
deposit,  which,  at  distances  of  two  car-lengths  apart,  is  connected 
with  the  surface  by  mill-holes  sunk  from  the  surface  or  from  the 
level  above  previously  worked  by  steam-shovels.  At  the  bottom 
of  the  mills  are  chutes  for  loading  into  cars  the  ore  blasted  down 
at  the  surface  and  delivered  through  the  mills,  to  be  thence  hauled 
out  of  the  tunnel. 

This  system  of  mining  is  very  successfully  loading  a  train  of 
twenty-five  2O-ton  cars  per  hour. 

Stripping  Anthracite  Coal. — A  lenticular  coal  basin  at  Latti- 
more,  Pennsylvania,  is  attacked  by  stripping  and  underground 
work.  The  depth  at  the  centre  of  the  deposit  is  about  100  feet. 
After  stripping  the  earth  cover  from  the  coal  a  slope  is  sunk 
on  the  foot-wall  of  the  basin  to  the  centre  or  lowest  point  of  the 
deposit.  Levels  are  driven  along  the  strike,  and  rooms  turned 
right  and  left,  which  are  worked  upward  to  the  surface,  the  coal 
being  allowed  to  fall  to  the  level  below.  Here  it  is  loaded,  trammed 
to  the  slope,  and  hoisted  out  of  the  mine.  The  product  of  the 
mine  is  limited  only  by  the  facilities  for  transportation. 

Underground  Mining. — The  deposits  having  a  tabular  form 
are  of  two  types:  beds,  or  seams,  occurring  nearly  horizontal, 
and  veins  whose  inclinations  approach  the  vertical.  The  former 
extend  laterally  over  a  considerable  territory  and  under  conditions 
which  are  quite  uniform  throughout,  permitting  a  system  of 
mining  to  be  promptly  selected.  Veins  present  so  many  eccen- 
tricities which  increase  with  the  progress  of  the  work  that  the 


METHODS  OF   MINING.  37 

early  operations  are  somewhat  tentative  until  the  property  has 
been  thoroughly  explored. 

Deposits  of  coal,  iron  ore,  and  minerals  occurring  in  the  stratified 
rocks  are,  with  rare  exceptions,  flat.  They  have  a  greater  regu- 
larity of  thickness,  inclination,  and  character  of  contents  than  do 
vein  deposits  in  the  metalliferous  regions  which  are  encased  in 
hard  rock,  occupy  a  vertical  position,  and  are  subject  to  changes 
not  only  in  dimensions  and  depth,  but  even  in  the  character  of 
their  mineral.  The  brittleness  of  the  two  minerals  is  markedly 
different.  Coal  must  be  delivered  in  large  sizes  for  the  market, 
while  metalliferous  ores  are  just  as  valuable  when  pulverulent  as 
in  the  lump.  The  entire  contents  of  the  coal-bed  can  be  utilized 
and  are  salable,  and  hence  are  extracted  as  completely  as 
possible.  On  the  other  hand,  in  the  precious  metal  deposits 
not  one  tenth  of  the  contents  between  the  walls  is  of  sufficient 
value  to  warrant  hoisting  to  the  surface;  the  remaining  nine 
tenths,  being  valueless,  must  be  stowed  away.  These  radical 
differences  in  the  character  of  the  deposits  require  distinct  systems 
of  extraction  for  coal  and  for  ore,  and  hence  they  will  be  treated 
separately. 

In  one  respect  only  is  the  mining  of  coal  and  the  mining  of 
other  minerals  conducted  alike.  The  operations  progress  to  the 
rise  or  upon  the  slope  except  on  exceedingly  rare  occasions, 
when  precious  metals  may  be  worked  downward. 

The-  thickness  and  pitch  of  the  mineral  deposit,  the  thickness 
and  nature  of  its  cover,  the  direction  and  its  outcrop,  the  mechani- 
cal and  physical  character  of  its  rock  or  mineral  having  thus 
been  ascertained,  the  place  of  opening  and  the  character  of  the 
entry  into  the  mine  are  selected  according  to  the  principles  dis- 
cussed in  the  previous  chapter.  The  directions  of  the  gang- 
ways to  be  used  for  haulage  and  of  the  rooms  for  economical 
work  are  next  determined  in  connection  with  the  system  to  be 
adopted.  After  this  the  vein  or  bed  is  blocked  out  and  is  ready 
for  mine  operations. 

Systems  of  Mining :  Beds. —  According  as  it  may  be  desirable 
or  necessary  to  support  the  stratum  above  the  mineral  bed  or 


38  MANUAL  OF  MINING. 

allow  it  to  fall,  the  mineral  should  be  removed  in  two  distinct 
operations,  or  should  be  extracted  in  its  entirety  at  once.  The 
first  is  the  pttlar-and-room  system,  in  which  chambers  or  rooms 
are  cut  out  at  various  points,  isolated  from  one  another  by 
untouched  portions  of  the  bed,  and  the  mineral  from  them 
extracted  on  the  advance  from  'the  shaft  or  entry  toward  the 
boundaries.  The  remainder  of  the  deposit  is  attacked  on  the 
return.  This  system  is  applicable  where  the  roof  is  sufficiently 
strong  to  be  left  with  a  partial  support. 

The  second,  longwatt  system,  contemplates  the  extraction  of 
the  entire  contents  of  the  bed  in  one  operation.  It  is  employed 
when  the  roof  is  treacherous  or  weak  and  cannot  be  economically 
secured.  No  effort  is  then  made  to  support  it.  By  this  plan 
the  entire  width  of  the  coal  property  is  attacked  simultaneously; 
or  mining  may  progress  circumferentially  and  outward  to  the 
boundaries,  removing  the  entire  mineral  contents;  or  all  coal 
may  be  removed  while  progressing  in  the  opposite  direction 
from  the  boundaries  to  the  shaft.  These  methods  are  known  as 
the  longwall  advancing  and  longwall  retreating  systems. 

Naturally  the  simplest  method  of  attacking  the  mineral 
would  be  to  complete  a  series  of  breasts  as  wide  as  the  nature 
of  the  roof  or  the  walls  would  permit,  and  opening  them  when 
the  main  headings  have  been  carried  to  the  extreme  boundaries, 
leaving  the  mineral  intact  except  for  the  necessary  roads.  The 
breasts  are  mined  out  as  rapidly  as  possible  on  the  return  home- 
ward toward  the  shaft.  This  plan  of  mining  retreating  represents 
the  maximum  degree  of  safety  for  miners,  who  are  then  working 
in  solid  mineral.  It  is  economical  for  operators,  who  may  recover 
the  entire  deposit.  The  cost  of  mine  timber  is  far  less,  and  the 
product  per  acre  greater,  than  by  mining  outward  from  the  shaft. 
But  the  plan  requires  patience  and  a  large  capital,  particularly 
if  the  area  to  be  mined  is  large  or  the  royalty  to  be  paid  is  great. 
It  cannot  be  unconditionally  recommended. 

Systems  of  Mining  Veins. — The  systems  employed  in  recovering 
mineral  from  veins  are  essentially  the  same  as  for  beds  with 
such  modifications  as  are  due  to  the  increase  in  pitch.  The 


METHODS  OF  MINING.  39 

rock  to  be  supported  is  the  vein  matter  itself,  while  the  rock  adjoin- 
ing the  vein  exerts  little  pressure  upon  the  vein  matter.  Hence 
the  character  and  strength  of  the  vein  matter  determines  the 
choice  of  the  system.  In  strong  mineral  and  hard  coal  the  pillar- 
and-room  system  is  adopted.  In  metalliferous  districts  the 
rooms  for  the  first  excavations  are  called  stopes,  and  the  system 
is  a  sloping  system.  The  longwall  system  as  applied  to  veins 
consists  in  the  attack  of  their  contents  over  a  broad  area  and 
allows  the  friable  materials  to  cave  without  restraint.  The  broken 
mineral  contents  are  then  removed  and  replaced  by  additional 
material  caved  from  above.  This  continues  until  the  lift  has  been 
exhausted.  This  caving  system  is  much  in  vogue  in  iron  mines. 

Two  modifications  bearing  resemblances  to  both  of  the  typical 
systems  are  also  in  extensive  use,  one  being  known  as  the  square- 
set  system  and  the  other  as  the  filling  system.  In  these  the  area 
of  attack  is  broad,  but  the  roof  is  not  allowed  to  cave.  The  exca- 
vations are  filled  as  rapidly  as  it  is  possible  to  supply  the  material 
for  support.  In  the  square-set  system  a  massive  cribwork  of 
timbers  is  built  into  position.  In  the  filling  system  foreign 
materials  are  lowered  into  the  mine  from  above,  spread  and 
packed  into  the  rooms  immediately  behind  the  miners  who  are 
engaged  in  extracting  mineral. 

Systems  of  Mining  Coal. — The  two  typical  systems  employed 
for  extraction  of  coal  are  the  pillar-and-room  and  the  longwall 
previously  described.  The  longwall  system  is  employed  for 
thin  seams  of  coal^  or  for  thick  seams  divided  into  thin  layers  by 
partings,  and  having  regular  thickness,  a  uniform  pitch,  weak 
roof,  soft  floor,  and  the  bed  deep.  The  inclination  of  the  beds 
for  the  longwall  system  must  not  exceed  10°. 

For  the  pillar-and-room  method  the  vein  may  be  thick,  the 
material  for  filling  and  gob  may  be  absent,  the  vein  shallow, 
the  coal  soft,  and  the  seams  irregular  in  thickness  and  eccentric 
in  pitch. 

As  not  all  of  the  conditions  favorable  for  either  system  exist 
in  a  given  mine,  the  choice  must  be  indicated  by  the  compara- 
tive merits  of  the  majority  of  favorable  conditions  of  greatest 


40  MANUAL  OF  MINING. 

importance  from  an  economical  standpoint.  No  system  can 
entirely  satisfy  all  the  conditions;  hence  a  compromise  inevitably 
results  in  adopting  the  least  objectionable  system,  with  such 
modifications  as  have  been  evolved  from  local  conditions  or  are 
based  upon  the  structural  difficulties  encountered  in  the  given 
mine.  The  widths  and  directions  of  the  rooms  or  the  stalls  for 
attack,*  the  mode  of  recovering  the  mineral  from  the  pillars,  and 
the  process  of  filling  and  packing  the  excavations  vary  in  locali- 
ties and  mines  and  cause  a  departure  from  the  typical  method 
in  some  cases  to  such  an  extent  that  the  identity  is  almost  lost. 

The  Longwall  Advancing  System. — This  system  consists  in 
driving  gangways  from  the  shaft,  leaving  around  the  latter  an 
ample  pillar  for  security  (Fig.  i).  When  its  boundaries  have 
been  reached  the  gangways  are  extended,  the  miners  proceeding 
to  the  right  and  left  from  them,  cutting  the  coal  while  advancing 
from  the  shaft  and  throwing  behind  them  such  waste  and  debris 
as  are  produced  during  the  operations.  No  effort  is  made  to 
support  the  roof  beyond  that  which  is  necessary  for  the  pro- 
tectipn  of  the  men  for  ten  yards  back  of  the  face.  The  gang- 
ways serving  as  roads  are  carried  with  the  progress  of  the  work 
by  building  pack-walls  out  of  the  material  obtained  from  the 
debris.  They  are  advanced  radially.  Each  length  of  working 
face  is  a  stall.  When  two  miners  are  assigned  to  a  stall  of 
about  30  feet  each,  the  gangways  or  haulage  roads  serving  two 
stalls  60  feet  long  are  not  over  60  feet  apart.  This  neces- 
sitates building  additional  roads  to  reduce  the  miners'  length  of 
haul. 

The  numerous  stalls  extend  in  one  continuous  straight  line, 
as  in  Figs,  i  and  4,  or  a  curved  line,  as  in  Fig.  2,  without 
any  corners  or  breaks.  The  roof  either  bends,  yields,  and  closes 
upon  the  excavation,  or  it  breaks  freely  and  fills  the  space  which 
is  known  as  the  gob.  The  subsidence  of  the  roof  is  communicated 
to  the  overlying  strata  and  will  often  extend  hundreds  of  feet 


*  The  reader  will  find  in  the  appendix  the  definitions  of  the  various  technical 
terms  employed  in  this  chapter  and  elsewhere  throughout  the  book. 


METHODS  OF  MINING. 


42  MANUAL  OF  MINING. 

toward  the  surface.    Hence  the  bed  must  be  deep  or  the  strata 
firm,  if  the  surface  land  is  not  to  be  injured. 

The  Scotch  System. — A  system  of  mining  known  as  the  45° 
system,  and  by  some  as  the  Scotch  system  of  longwall,  is  shown 
in  Fig.  2.  The  main  entries  aa  are  at  right  angles  to  each  other, 
being  turned  from  the  right  and  left  from  the  upcast  shaft  c.  The 
gangways  bb  are  turned  at  an  angle  of  45°  from  the  main  entries. 


FlG.  2.— The  Scotch  Longwall  System. 

The  shaft  pillars  p  are  left  untouched ;  the  dimensions  of  each  are 
discussed  in  Part  II,  Chapter  I.  The  face  of  the  main  entry  is 
advanced  with  a  width  of  30  feet,  allowing  12  feet  between  the 
pack -walls,  or  9  feet  on  each  side.  Miners  turning  from  right  or 
left  of  the  head  of  the  entry  break  the  coal  over  its  entire  face  and 
advance  outward,  shipping  the  mineral  through  the  roads.  The 
pack-walls  are  extended  on  either  side  as  fast  as  the  face  pro- 
gresses. The  roads  are  so  driven  as  to  leave  from  120  to  180 
feet  of  face,  according  to  the  nature  of  the  room. 

Cutting  the  Coal. — The  process  of  mining  the  coal  is  shown  in 
Fig.  3  at  a,  where  a  groove  is  cut  in  the  floor  or  near  the  floor 


METHODS   OF  MINING.  43 

by  pick  or  by  machine.  Two  rows  of  props  bb  are  placed  parallel 
to  the  face  and  at  such  distances  from  it  as  to  give  ample  support 
without  interfering  with  the  work  of  the  miner.  The  sprag  c 
supports  the  corner  of  the  underhold  coal  while  the  miners  are  at 
work.  The  sprags  are  removed,  allowing  the  roof  to  release  the 
coal  as  soon  as  the  length  of  the  stall  has  been  undermined.  No 
powder  is  used,  the  pressure  of  the  roof  being  employed  for  the 


FIG.  3. — A  Cross-section  at  the  Longwall  Face. 

purpose.  The  slate  and  debris  are  thrown  on  the  gob  behind 
the  men;  the  back  line  of  props  is  withdrawn  and  advanced  in 
front;  the  track  is  advanced  in  sections  parallel  to  the  face;  and 
the  operation  is  repeated  for  another  3  to  5  feet  according  to 
the  depth  of  the  groove.  If  the  roof  is  weak,  little  space  is  left 
for  the  track  and  the  gob-roads  are  nearer  together,  the  stall 
being  shorter. 

When  the  inclination  of  the  coal-seam  exceeds  10°  the  com- 
ponent of  the  vertical  pressure  of  the  overlying  strata  transmitted 
by  the  roof  to  the  coal  parallel  to  the  bed  of  the  vein  causes  a 
side  pressure  or  "swing"  which  tends  to  produce  slip  in  the  coal. 
This  is  serious,  and  if  the  curved  face  of  the  stall  will  not  provide 
for  it  the  system  must  be  abandoned. 

Usually  the  mining  is  conducted  by  day,  while  the  construction 
of  the  pack- walls  and  the  driving  of  gob-roads  proceeds,  at  night. 
If  it  is  necessary  to  operate  in  two  shifts,  the  stalls  are  worked 
alternately  from  the  main  entries. 

The  Demerits  of  the  Longwall  System.— The  success  of  the  sys- 
tem depends  upon  the  nature  of  the  roof  and  its  behavior  when 
the  coal  is  removed.  It  should  be  brittle  and  weak.  A  tougb 


44  MANUAL  OF  MINING. 

flexible  roof  is  unsuitable  and  may  require  blasting  down  to 
relieve  the  pressure  upon  the  coal.  The  roads  which  are  main- 
tained through  the  excavations  are  called  the  gob-roads,  and  here- 
in lies  the  serious  objection  to  this  system.  The  fracturing  of 
the  roof  which  is  the  essential  element  of  success  tends  also  to 
close  the  roads,  increase  the  expense  of  mining,  and  endanger  the 
lives  of  the  men.  The  roof  and  floor  are  continually  being  cut  into 
to  maintain  sufficient  height  for  the  travel  in  the  roads.  The  length 
and  number  of  the  latter  increase  with  the  progress  of  the  work, 
and  the  cost  of  their  maintenance  becomes  a  serious  one.  If 
the  boundary  is  a  great  distance  away,  the  expense  of  keeping 
them  open  may  be  so  great  as  to  necessitate  a  discontinuance  of 
the  system.  The  liability  of  the  material  in  the  gob  to  spon- 
taneous combustion  and  fire  is  an  additional  serious  danger.  A 
fire  once  begun  cannot  be  extinguished  without  the  entire  removal 
of  the  waste.  It  is  impossible  to  make  the  pack-walls  air-tight. 
Ventilation  presents  difficulties,  for  the  numerous  roads  must 
receive  separate  air-currents,  while  the  working  face  may  easily 
be  ventilated  by  a  single  continuous  current.  • 

Longwall  Retreating. — This  is  considered  a  better  and  safer 
system,  as  the  roof  pressure  is  supported  by  the  solid  coal  while 
the  work  is  proceeding  toward  the  shaft.  The  waste  produced 
in  mining  is  allowed  to  accumulate  behind  the  men  where  the 
roof  closes  upon  the  gob.  There  is  no  danger  of  any  section  of 
the  mine  being  closed  by  collapse,  nor  is  there  any  fear  of  spon- 
taneous combustion. 

In  Fig.  4  is  shown  a  system  of  opening  the  mine  by  longwall 
retreating.  The  four  pairs  of  entries  aa  are  started  from  the 
downcast  shaft  and  driven  at  right  angles  to  one  another  to  the 
boundaries  between  them;  entry  pillars  cc  are  left  about  40  feet 
thick  with  breakthroughs  at  intervals  of  100  feet  apart.  When 
the  entries  have  reached  within  300  feet  of  the  boundary  gangways 
dd  are  driven,  and  at  right  angles  to  them  the  face  entries  ee,  at 
intervals  of  200  or  300  feet.  From  their  heads  the  stalls  are  turned 
off  to  the  right  and  left,  opening  a  continuous  face  of  coal.  Mining 
is  then  commenced.  The  rails  are  laid  and  the  haulage-ways 


METHODS  OF  MINING.  45 

mm  moved  from  the  entries  ee,  where  they  are  no  longer  of  service. 
Work  is  simultaneously  commenced  in  the  haulage-ways  nn  to 
prepare  for  the  continuance  of  the  retreat  from  the  gangways 
dd  as  soon  as  the  operations  have  been  completed  on  the  face 
entries  ee. 

The  main  haulage-ways  are  in  the  return  airways  to  the 
upcast  and  hoisting  shaft  u.  Entries  crossing  the  intake  air- 
ways are  carried  over  by  air-bridges,  Chapter  XIV. 


FIG.  4. — Longwall  System  of  Mining  Retreating. 

Occasionally,  when  the  roof  is  very  weak,  the  gangways  are 
driven  singly  and  connected  by  cross-cuts  60  feet  apart  instead 
of  being  driven  in  pairs.  This  method,  however,  is  productive 
of  a  greater  amount  of  narrow  work  and  a  greater  expense  in 
ventilation. 

The  Combination  Longwall  System. — In  Fig.  5  is  illustrated 
a  combination  of  the  retreating  and  advancing  systems.  A 
portion  of  the  bed  above  the  bottom  of  the  shaft  is  worked  by  the 
advancing  system,  while  that  below  is  driven  by  the  retreating 
system.  The  seam  is  16  feet  thick,  but -does  not  prove  very  profit- 


46  MANUAL  OF  MINING. 

able  because  of  a  considerable  expense  incurred  in  driving  the 
entries  while  planning  the  works.  However,  when  properly  ex- 
ecuted it  will  relieve  the  roof,  though  there  is  always  danger  of 
a  suspension  of  its  subsidence  which  interferes  with  the  "  swing, " 
and  also  in  a  collapse  along  an  unexpected  line. 

The  mining  to  the  rise  is  conducted  advancing  in  section  A, 
while  in  section  B  the  operations  are  in  retreat. 


FlG.  5. — A  Combination  of  Retreating  and  Advancing  Longwall  Systems. 

Comparative  Merits  of  Longwall  Retreating  and  Longwall 
Advancing. — In  the  matter  of  ventilation  the  retreating  system 
is  more  perfect  than  the  advancing  system,  as  it  can  be  regulated 
to  better  advantage.  No  danger  need  be  apprehended  from  the 
evolution  of  gas  from  the  waste.  The  cost  of  mining  is  the  same 
in  both  systems,  but  the  expense  of  making  pack-walls  and  of 
opening  and  maintaining  the  roads  is  far  greater  than  the  in- 
creased cost  of  driving  the  narrow  gangways  through  the  solid 
coal.  The  large  initial  outlay  in  longwall  retreating  for  the  four 
pairs  of  main  entries  out  to  the  extreme  boundary  of  the  property, 
and  for  laying  tracks  the  entire  length,  and  the  long  period  of 


METHODS  OF  MINING.  47 

waiting  for  the  returns,  constitute  the  chief  obstacle  to  its  more 
ready  acceptance.  But  its  aggregate  advantages  warrant  its 
employment  under  all  circumstances  where  the  longwall  system  is 
at  all  applicable. 

The  Nottingham  or  Barry  Modification  is  employed  for 
mining  splint  coal  of  five  feet  in  thickness  with  a  strong  roof  and  a 
clay  floor.  In  this  the  gangways  are  6  feet  wide  in  the  solid  coal, 
with  2o-foot  pillars  between  them  and  the  cross-entries  200  feet 
apart.  From  these  are  driven  flats  120  feet  apart  and  200  feet 
long,  the  coal  being  attacked  "long  horn,"  Fig.  9,  and  to  the  rise. 
The  cars  travel  along  the  working-face.  Eight  miners  are  em- 
ployed on  each  stall  of  120  feet,  and  produce  large  coal  at  low 
cost,  with  an  excellent  yield  per  acre. 

In  one  mine,  changing  from  room-and-pillar  to  longwall,  the 
coal  was  opened  by  butt-entries  on  both  sides  of  the  gangways  until 
their  faces  met,  the  stalls  being  48  feet  wide.  The  coal  was 
strong  and  had  a  regular  cleat  overlaid  by  a  slate  top  with  a  hard 
floor.  It  was  blasted,  not  undercut,  and  37  tons  per  keg  of 
powder  were  obtained. 

Pillar-and-room  System. — The  second  system  of  mining  coal 
consists  in  driving  levels  and  opening  long  narrow  rooms,  from 
which  the  coal  is  extracted  while  advancing  from  the  shaft,  and 
in  leaving  pillars  on  either  side  of  untouched  coal  which  are  sub- 
sequently recovered  on  the  return.  The  rooms  are  frequently 
termed  stalls  and  breasts,  though  the  latter  term  is  more  cor- 
rectly applied  to  the  face  of  attack.  When  the  rooms  are  very 
broad  compared  with  their  length,  the  term  chambers  is  some- 
times applied  to  designate  them.  The  pillars  left  for  support 
are  also  called  ribs. 

This  system  is  a  survival  of  antiquated  methods  and  still  pre- 
vails to  a  greater  extent  in  the  United  States  than  elsewhere, 
rather  because  of  local  custom  than  from  any  special  merit 
which  it  possesses.  It  is  employed  where  the  material  for  filling 
or  gob  is  scarce.  A  great  diversity  exists  in  the  various  localities 
in  the  details  of  the  plan.  The  height  of  the  lifts,  the  direction 
of  the  rooms,  their  dimensions,  the  forming  of  pillars,  the  rob- 


48  MANUAL  OF  MINING. 

bing  of  pillars,  the  taking  down  of  roof,  etc.,  vary  widely. 
Many  of  its  modifications  are  due  to  physical  conditions  of  the 
coal  and  roof,  though  the  latter  elements  do  not  always  receive 
the  consideration  due  them.  The  facility  with  which  the  details 
of  working  may  be  varied  to  suit  the  irregularities  of  pitch 
and  thickness,  the  rolls  in  the  strata,  etc.,  renders  it  especially 
acceptable  in  regions  of  geological  disturbances. 

The  necessity  for  economy  in  the  conservation  of  the  mineral 
is  not  imposed  upon  the  American  operator.  The  question  of 
output  per  acre  is  not  of  vital  importance,  and  mining  as  a  con- 
sequence is  conducted  here  in  rather  a  loose  manner.  Indeed, 
with  the  exception  of  culm-flushing  there  has  been  no  improve- 
ment  introduced  in  the  anthracite  exploitation  during  the  past 
sixty  years,  though  numerous  mechanical  improvements  have  been 
utilized  to  economize  the  handling  and  preparation  of  the  coal. 

The  Typical  Room-and-pillar  System. — As  stated  in  the  pre- 
vious chapter,  the  height  of  the  lifts  varies  with  the  thickness  and 
inclination  of  the  seams.  But  this  determined,  the  gangways 
are  driven,  after  which  the  rooms  are  turned  as  fast  as  practicable 
to  the  rise  or  up  the  slope  at  an  angle  depending  upon  the  pitch, 
as  in  Figs.  6  and  7,  or  on  the  mode  of  conveying  the  coal  from 
the  breasts  to  the  gangways.  The  jaws  or  necks,  aa,  of  the 
rooms  B  are  about  6  feet  wide  for  a  distance  equal  to  that  allowed 
for  the  stump-pillars  A,  when  they  are  suddenly  enlarged  to  a 
working-face  of  20  or  30  feet,  as  time  and  the  condition  of  the 
roof  will  allow.  The  rooms  are  driven  regular  and  uniform 
to  a  length  of  from  eight  to  ten  times  their  width,  leaving  a 
chain-pillar  for  support  to  the  upper  level.  The  rooms  of  each 
lift  are  opened  progressively  to  the  boundary  as  fast  as  circum- 
stances will  permit. 

The  cost  of-  driving  narrow  entries  and  headings  aa  and  cc 
is  so  high  that  modifications  are  introduced  whereby  the  jaws  aa 
are  made  as  short  as  possible  and  widened  out  to  the  full  face  of 
the  room,  thus  reducing  the  narrow  work  without  weakening  the 
stump-pillars.  The  pillars  P  left  between  the  rooms  are  unbroken 
except  for  three  or  four  small  breakthroughs  dd,  through  which 


METHODS  OF  MINING.  4^ 

the  return  ventilating  air-current  passes  from  room  to  room  or* 
its  way  to  the  upcast. 

Rock  Pressure. — Assuming  the  specific  gravity  of  rock  to  be- 


2.6,  then  a  block  having  a  base  of  i  square  inch  and  a  height 
of  12  inches  will  weigh  i  pound.  A  column  500  feet  high  will 
exert  a  pressure  upon  its  base  of  500  pounds  per  square  inch. 
Each  square  foot  of  rock  at  that  depth  below  the  surface  will. 


MANUAL  OF  MINING 


I 


METHODS   OF   MINING  51 

receive  a  pressure  of  72,000  pounds  per  square  inch.  The  load 
which  a  pillar  of  coal  will  sustain  is  therefore  proportional  to 
its  depth  below  the  surface  and  the  area  of  its  base.  A  coal- 
pillar  25  feet  wide  and  200  feet  long  at  a  depth  of  500  feet  will 
receive  a  total  pressure  of  180,000  tons.  To  sustain  this  load 
without  suffering  injury  it  must  have  a  strength  at  least  equal 
to  two  or  three  times  the  load  thus  imposed  upon  it. 

The  Strength  of  Coal. — Experiments  recently  conducted  upon 
the  crushing  strength  of  coal  have  obtained  values  ranging 
between  1600  pounds  per  square  inch  and  2100  pounds  per  square 
inch.  The  value  for  the  ultimate  strength  of  coal  having  been 
determined  in  a  specific  case,  it  is  possible  to  determine  the 
dimensions  of  a  pillar  required  to  sustain  a  given  load  trans- 
mitted to  it  from  the  roof. 

Owing  to  the  wide  range  of  values  obtained  for  the  crushing 
strength  of  coal,  it  is  advisable  that  each  engineer  determine  the 
specific  value  for  the  coal  of  his  locality,  determining,  thereupon,, 
the  dimensions  of  the  pillars  to  be  used  in  connection  with  this 
system  of  mining.  Sample  cubes  should  be  selected  and  sawed 
from  various  portions  of  the  seam,  and  their  strength  normal  to 
the  bedding  ascertained  by  actual  test.  They  should  be  pro- 
cured from  the  breasts  rather  than  the  pillars. 

Influence  of  Cleats  upon  the  Resistance  of  Coal. — It  wa& 
shown  in  the  previous  chapter  that  the  direction  of  the  cleavage 
of  coal  may  determine  the  direction  of  driving  the  rooms  and' 
roads.  The  cleavage  may  also  determine  the  resisting  strength 
of  the  coal.  In  Fig.  8  is  illustrated  a  block  of  coal  with  its  hori- 
zontal bedding  cleavage  aa,  and,  extending  from  roof  to  floor,, 
the  vertical  cleavage  bb,  as  well  as  the  butt-cleats  cc,  at  right 
angles  to  both.  In  all  coals  this  prismatic  structure  is  pronounced 
in  one  or  more  directions,  and  along  these  planes  a  slip  may  occur 
before  actual  crushing  ensues.  The  load  which  the  coal  can 
sustain  is  therefore  less  than  its  actual  strength. 

In  Fig.  9  are  shown  arrows  indicating  the  directions  of  attack 
upon  the  lines  representing  the  plane  of  the  main  cleavage  bb. 
The  butt-cleat,  line  cc,  presents  usually  a  hackly  appearance 


MANUAL  OF  MINING. 


when  exposed,  and  can  be  readily  distinguished  from  the  main 
cleats,  which  are  perfectly  smooth.  Adhesion  along  the  plane  of 
the  former  is  stronger,  and  hence,  when  conditions  of  haulage  do 


FIG.  8.— The  Cleats  in  Coal. 


FIG.  9. — Directions  of  Attack  upon 
the  Cleats. 


not  prevent,  most  mining  is  conducted  in  the  direction  represented 
l>y  the  arrow  c  at  right  angles  to  the  main  cleat,  known  as  "  face- 
on  work";  an  advance  in  the  direction  of  the  arrow  e  is  "end-on 
work"  mining;  and  in  the  direction  of  arrow  d,  "half  end-on 


FIG.  12.  c       c> 


The  Directions  of  Cleavage. 

work."  The  mining  of  coal  at  an  angle  greater  than  45°  and 
less  than  90°  is  said  to  be  "long  horn,"  arrow  g;  and  at  an 
angle  of  less  than  45°,  arrow  /,  is  called  "short  horn." 

In  Fig.  10  is  illustrated  the  columnar  structure  of  coal.  Across 
the  bedding  a  it  is  stronger  than  the  cleat  &,  represented  by  the 
broken  lines.  In  the  horizontal  position  a  is  weaker  than  b  (Fig. 
IT).  In  the  inclined  position  (Fig.  12)  neither  cleavage  plane 
is  in  its  original  position.  The  cleavages  receive  their  respective 
components  of  the  thrust  c,  the  component  a  acting  normal  to  the 


METHODS  OF  MINING.  S3 

bedding  planes  or  across  the  main  cleat,  while  the  component 
e  is  parallel  to  the  main  cleat.  The  amount  of  thrust  falling 
normally  upon  the  bedding  plane  is  less  as  the  inclination  increases. 

Size  of  Coal-pillars. — The  width  of  the  coal-pillar  depends 
upon  the  strength  of  the  coal  and  the  depth  of  the  rock,  rather 
than  on  the  strength  of  the  roof.  When  the  rooms  are  open  each 
pillar  must  support  not  only  the  rock  immediately  over  it,  but  also 
sustain  the  pressure  transmitted  to  it  from  the  roof  over  the  room. 
A  rigid  and  firm  roof  will  transmit  its  entire  pressure  equally  to 
the  sides.  Hence  the  coal,  at  a  depth  of  500  feet,  in  the  pillars  25 
feet  wide  and  200  feet  long  between  rooms  20  feet  wide,  will 
receive  a  pressure  equal  to  the  weight  of  a  mass  35  feet  wide,  200 
feet  long,  and  500  feet  high,  or  about  252,000  tons.  This  corre- 
sponds to  50.4  tons  per  square  foot.  So  long  as  this  is  less  than 
the  breaking  strength  of  the  coal  the  pillar  will  remain  intact.  If 
the  crushing  strength  be  2000  pounds  per  square  inch,  the  stress 
upon  the  coal  is  less  than  half  its  ultimate  strength.  The  margin 
or  factor  of  safety  is  2.8.  This  factor  will  be  further  reduced  if 
the  planes  of  main  cleavage  are  pronounced  and  weak. 

At  a  depth  of  1400  feet  there  remains  no  margin  for  safety, 
the  limit  of  the  coal  resistance  being  reached  with  its  relation  of 
room  to  pillar.  Below  this  depth  the  pillars  must  be  wider  or  the 
rooms  narrower  if  the  system  is  to  be  pursued.  When  a  roof  is 
brittle  it  will  break  without  transmitting  its  pressure  to  the  pillar. 
The  pillar  is  relieved  of  some  pressure,  in  which  event  its  width 
is  not  necessarily  increased.  A  soft  floor  requires  a  broad  pillar 
to  distribute  the  thrust  over  a  large  area  of  the  floor.  Otherwise 
a  heaving  of  the  floor  ensues,  which  disturbs  the  tracks  and 
" creeps"  over  the  territory,  closing  rooms  and  airways. 

The  Relative  Dimensions  of  Pillars  and  Rooms. — The  width 
of  the  pillars  depends  upon  the  character  and  length  of  service 
they  are  to  perform.  Safety  demands  large  pillars,  but  inasmuch 
as  only  a  portion  of  their  coal  contents  is  recoverable  during  the 
second  working,  a  very  large  percentage  of  the  total  in  the  tract  is 
lost.  A  rapid  production  demands  wide  rooms  which  can  be 
mined  by  machinery  and  exhausted  quickly,  but  they  require  a^ 


54  MANUAL   OF  MINING. 

good  roof  and,  in  turn,  broad  pillars  to  support  them.  The  rela- 
tion between  the  two  must  be  determined  by  local  conditions, 
observing  the  principles  mentioned  above.  Narrower  pillars  with 
wide  rooms  are  decidedly  wasteful  of  coal,  and  in  a  five-foot  seam 
with  ribs  of  12  feet  and  rooms  36  feet,  none  of  the  pillar  coal  is 
recovered,  whereas  in  a  6-foot  seam  with  1 2-foot  rooms  and 
28- foot  pillars  nearly  90  per  cent  of  the  coal  contents  are  sold, 
particularly  if  panel  pillars  are  also  employed  in  sections  of  the 
mine,  as  illustrated  in  Fig.  17.  The  usual  proportion  of  pillar  to 
room  is  about  3  to  2  for  the  average  depth  of  mining  as  pre- 
vailing at  present  in  America. 

As  an  illustration,  it  will  be  found  on  calculation  that  in  a 
mine  opened  with  rooms  of  12  feet  width  between  pillars  of  18 
feet  a  greater  concentration  of  labor  is  obtained,  and  there  are 
positive  sources  of  economy  over  that  possible  if  the  rooms  and 
pillars  alike  are  20  feet  wide.  Along  a  mile  of  gangway  176 
rooms  can  be  opened  in  the  former  and  but  132  rooms  in  the 
latter  design.  Though  the  aggregate  breast  length  is  less,  there 
will  be  more  rooms  for  simultaneous  work.  Or,  for  an  equal 
number  of  working-faces,  there  will  be  a  concentration  of  work 
by  the  narrower  pillars  and  rooms,  which  saves  3520  feet  of  track 
for  the  132  rooms;  haulage  will  be  less;  and  44  cars  can  be  ob- 
tained in  the  same  time  from  three  quarters  the  distance  of  that 
when  the  rooms  are  spaced  40  feet  apart.  In  robbing  the  pillars 
more  coal  will  also  be  recovered  from  the  relatively  thicker  pillars. 

The  width  of  pillars  is  increased  and  that  of  the  rooms  decreased 
as  the  depth  of  the  vein  is  great.  The  limit  to  an  economical 
application  of  this  system  is  reached  when  the  rooms  are  reduced 
to  an  unprofitable  portion  of  the  deposit.  At  a  depth  of-  700 
feet  they  contain  less  than  30  per  cent  of  the  total  contents  of 
the  vein.  At  1600  feet  depth  the  pillars  leave  only  25  per  cent  of 
the  coal  for  the  rooms;  while  at  2000  feet. the  rooms  are  one 
fourth  as  wide  as  the  pillars,  and  are  capable  of  producing  not 
more  than  15  per  cent  of  the  total  coal. 

The  general  impression  that  coal-pillars  and  beds  of  steep 
pitch  are  less  strong  than  those  in  flat  beds  is  untrue  unless  a 


METHODS   OF  MINING.  55 

deformation  of  the  strata  has  produced  a  fissility  in  it.  Their 
resistances  are  equally  great  and  their  life  as  great,  being  dimin- 
ished only  by  a  protracted  period  of  pressure  and  by  alterations  of 
temperature  and  humidity  which  injure  the  integrity  of  the  pillar. 
This  is  independent  of  the  inclination  of  the  bed. 

There  is  an  economical  limit  to  the  narrowness  of  the  rooms, 
based  not  upon  physical  laws  but  upon  an  artificial  law  by 
which  a  working- face  of  less  than  15  feet  is  considered  in  many 
coal-fields  as  "narrow  work,"  making  the  cost  of  driving  such 
rooms  greater  per  foot  of  length  than  that  of  wider  rooms.  More- 
over, narrow  rooms  do  not  admit  of  an  improvement  of  machine- 
cutters  or  large  gangs  of  loaders,  and  therefore  increase  the  cost 
of  coal. 

The  Mode  of  Delivering  the  Product  from  the  Face  to  the 
Entry. — The  most  important  problem  in  mining  is  the  method 
of  transporting  the  mineral  from  the  working-face  to  the  entry  or 
gangway.  It  usually  determines  the  plan  of  operations.  When 
the  vein  is  vertical  or  nearly  so  the  mineral  may  be  dropped 
through  mill-holes  or  chutes  to  the  cars  in  the  level  below  (Fig.  87), 
or  lowered  by  means  of  a  self-acting  incline  (Fig.  115).  In  the 
anthracite  region  the  working-face  is  termed  a  breast.  Here  the 
broken  coal  is  allowed  to  accumulate  in  the  batteries  (Fig.  13), 
which  are  loaded  at  the  coal  face  and  more  or  less  completely 
emptied  at  the  foot  from  a  platform,  the  flow  of  mineral  being 
controlled  by  a  gate.  So  long  as  the  pitch  exceeds  the  natural 
slope  of  broken  rock,  40°,  the  coal  will  roll  or  slide  on  the  rock 
bottom  without  the  necessity  for  lining  the  latter.  The  sides 
may  be  built  of  plank  spiked  to  the  props,  reaching  diagonally 
from  the  floor  to  the  roof,  and  the  lower  portion  closed  by  a 
gate  or  door  with  a  platform.  By  keeping  the  chute  or  battery 
full  of  coal  and  emptying  only  the  excess  through  the  gate,  the 
loss  from  attrition  will  be  slight.  This,  of  course,  is  not  pos- 
sible with  bituminous  coal.  At  a  grade  flatter  than  40°  the 
chute  must  be  floored  with  wood  or  iron,  but  below  18°  soft 
coal  will  neither  slide  nor  roll.  If  the  working-room  is  large, 
the  breasts  wide,  or  the  vein  thick,  the  chutes  are  carried  up  with 


56  MANUAL  OF  MINING. 

the  work,  one  on  each  side  of  the  waste  filling.  In  small  rooms 
the  chutes  are  built  centrally  with  the  manway  compartment 
alongside  of  them  (Fig.  13). 

Beds  having  an  inclination  between  18°  and  6°  from  the  hori- 
zontal are  the  most  difficult  to  provide  with  an  economical  method 
of  transportation.  They  are  too  flat  for  self-acting  systems  and 
too  steep  for  any  simple  plan  of  haulage.  An  endless- rope  way 
plan  for  each  room  is  too  complex,  and  any  scheme  involving 
shovelling  would  be  both  expensive  and  injurious  to  the  coal. 


FlG.  13. — Battery  Mining  of  Thick  Steep  Seams. 

In  such  event  buggy- roads  are  resorted  to  for  the  rooms  (Fig.  6). 
The  track  is  laid  in  sections  directly  upon  the  floor,  or  upon 
the  trestle  raised  at  the  lower  end  to  moderate  the  grade,  accord- 
ing to  the  headroom  that  may  be  afforded  and  the  pitch  of  the 
seam.  Another  method  consists  in  having  jig-planes  where  the 
loaded  car  in  its  descent  pulls  up  a  weight  which,  on  the  next 
down  trip,  raises  the  empty  car  back  to  the  face. 

Breasts  driven  up  a  grade  of  less  than  10°  are  said  to  be  flat, 
and  the  accessory  haulage  from  them  presents  no  special  diffi- 
culty. The  rooms  may  be  driven  directly  up  the  rise  for  the 
small  angles,  or,  when  the  pitch  reaches  the  limit,  are  laid  so 
as  to  secure  a  satisfactory  grade.  In  such  cases  mules  may  be 
utilized  for  the  power,  or  in  exceptional  cases  man-power  is 
employed. 


METHODS  OF  MINING.  57 

Rooms  in  beds  as  flat  as  3°  are  driven  directly  up  on  the  pitch, 
employing  mules  or  men  for  transportation  of  the  cars. 

The  rooms  are  rarely  worked  downward  in  the  dip,  unless  the 
pitch  be  very  slight.  For  a  short  distance  an  undulating  seam 
may  require  this  modification. 

Robbing  the  Pillars. — The  rooms,  having  been  carried  to  their 
length,  are  then  turned  and  "butted  off,"  the  pillar  at  the  top 
driving  along  the  under  side  of  the  chain-pillar.  The  pillars 
are  then  attacked  and  their  contents  removed  as  soon  as  possible, 
before  their  coal  contents  have  suffered  in  quality  from  pressure. 
The  operation  is  performed  as  rapidly  as  possible  for  safety,  and 
the  pillars  are  taken  in  successive  retreating.  Their  sides  are 
scaled  off  as  much  as  can  be  secured  before  the  pillar  crushes. 
The  heading  may  be  driven  off  the  centre  of  the  pillar  and  some 
of  the  middle  portion  of  the  coal  obtained,  leaving  two  thin  sup- 
ports for  the  roof;  or  the  pillar  may  be  worked  over  its  entire 
width  from  the  chain-pillar  toward  the  gangway. 

The  stump-  and  the  chain-pillars  are  not  disturbed  until  the 
lift  which  they  support  is  to  be  abandoned,  when  their  contents 
are  promptly  extracted  and  the  roof  allowed  to  cave. 

Fig.  14  is  the  plan  of  a  flat  seam  operated  as  described 
above,  the  two  sections  of  he  property  being  separated  by  the 
double  broken  lines.  From  the  outcrop  at  a  and  b  the  main 
and  return  headings  are  carried  60  feet  apart.  The  cross-entries 
are  360  feet  apart  and  the  rooms  300  feet  long.  The  outer  irreg- 
ular line  is  the  outcrop.  The  fan  is  located  at  c.  The  break- 
throughs d  are  midway  in  the  rooms  and  on  one  line  to  serve 
as  an  auxiliary  haulage-way  during  the  robbing  of  the  pillars. 
The  jaws  of  the  rooms  are  10  feet  wide  and  15  to  20  feet  long. 
The  room  is  widened  in  a  direction  away  from  the  outcrop. 
Pillar-robbing  begins  at  the  outcrop.  This  process,  being  slow 
and  decayed  till  all  rooms  are  mined,  results  in  an  excessive 
waste  of  coal.  The  speedy  returns,  good  ventilation,  and  safe 
work  commend  it. 

Comparison  of  Longwall  Advancing  with  Pillar-and  room 
System. — The  relative  advantages  and  disadvantages  of  these 


58  MANUAL   OF   MINING. 

two  systems  are  easily  stated.  In  the  former  the  expense  of 
maintaining  roadways  is  higher,  the  number  of  accidents 
greater,  the  product  of  round  coal  much  larger,  the  product 


JJJJJJJJJJJJJJJ 


FlG.  14. — A  Double-entry  System  in  a  Flat  Seam. 

per  acre  higher,  the  ventilation  of  the  face  simpler,  the  amount 
of  narrow  work  less,  and  the  consumption  of  powder  less.  On 
the  other  hand,  the  latter  system  is  more  advantageously  em- 
ployed where  faults,  horsebacks,  and  dikes  may  be  encountered, 


METHODS   OF  MINING. 


59 


if  the  upper  strata  are  wet,  if  the  boundary  is  at  a  great  distance, 
if  the  mines  are  shallow,  or  if  the  surface  land  is  very  valuable. 
Modified  Pillar-and-room  Systems.  —  In  some  fields  the 
cross-entries  are  driven  360  feet  apart  and  the  rooms  are  turned 
from  each  entry,  making  them  about  180  feet  long;  in  such  event 
the  cross-entries  are  carried  up  the  slope  independent  of  the 
direction  of  the  cleavage.  The  only  feature  commending  this 
plan  is  the  reduction  in  the  amount  of  narrow  work  and  the 
yardage  thus  saved. 


FACE  HEADINC 


BUTT  ENTRIE 
FlG.  15. — Connellsville  Method. 

A  modified  double-entry  system  employed  in  the  Connellsville 
coke  district  drives  gangways  1000  feet  apart  (Fig.  15),  and 
four  pairs  of  headings  from  them  at  distances  of  400  feet.  The 
rooms  are  turned  when  the  latter  entries  have  gone  1040  feet  and 
are  butted  off  at  the  chain-pillars,  after  which  the  pillars  are 
robbed.  The  coal  being  soft,  the  headings  and  the  jaws  of  the 


6o 


MANUAL  OF  MINING. 


rooms  are  driven  8  feet  wide,  the  pillars  30  feet,  and  the  stump- 
pillars  between  gangways  and  headings  52  feet. 

The  double-room  modification  giving  satisfaction  in  the 
Southern  States  for  seams  of  5  feet  thick  with  weak  roof  is  shown 
in  Fig.  16.  The  main  and  cross  entries  are  driven  in  pairs,  and 
later  the  rooms  are  started  right  and  left  42  feet  wide  with  pillars 
of  36  feet;  their  length  is  180  feet,  but  they  are  not  opened  fully 
until  the  cross-entries  have  reached  their  boundaries.  Props 


FIG.  16. — The  Double-room  System. 


are  set  along  the  inner  side  of  each  roadway  for  the  gob,  and 
waste  is  thrown  into  the  centre  of  the  room.  When  they  have 
reached  their  extreme  length  the  rooms  are  widened  and  one 
half  of  the  pillar  on  the  floor  withdrawn  retreating.  When  the 
succeeding  rooms  have  reached  their  limit  each  is  widened  in 
turn,  the  remaining  half  pillar  on  one  side  and  one  half  the 
adjoining  pillar  on  the  other  side  are  cut  away.  Props  set  along 
on  either  side  of  each  roadway  are  removed  or  broken  during 
the  retreat,  in  order  to  facilitate  the  uniform  fall  of  the  roof. 


METHODS   OF   MINING.  6 1 

Mining  Steep  Thick  Seams  of  Coal. — The  coal  in  a  steep  bed 
may  be  attacked  in  steps  proceeding  upward  to  the  top  of  the 
given  lift,  the  excavated  space  being  filled  with  rock  obtained 
from  elsewhere.  In  such  an  event  provision  must  be  made  for 
planked  chutes  to  deliver  the  coal  without  loss  or  admixture 
with  the  waste  rock.  The  broken  coal  may  be  allowed  to  accu- 
mulate in  the  excavated  spaces  as  a  partial  support  to  the  roof, 
and  also  for  the  men  while  mining  the  coal  overhead.  If  the 
rooms  are  not  too  wide,  the  coal  will  not  suffer  in  quality  by  the 
time  the  room  is,  exhausted.  The  other  system  of  mining  thick 
seams  of  coal  contemplates  its  removal  over  its  entire  length, 
proceeding  downward  from  the  top  of  the  lift  or  from  the  super- 
ficial covering  of  the  seam  and  allowing  the  material  above  to 
cave  upon  the  work  as  it  progresses  downward.  These  methods 
resemble  somewhat  the  two  longwall  systems  as  applied  to  the 
flat  thin  beds. 

The  famous  mammoth  bed  of  the  anthracite  fields  of  Penn- 
sylvania presents  the  greatest  difficulties  to  the  operator.  It 
varies  in  pitch  between  wide  limits  and  within  a  small  area 
attains  to  a  thickness  of  no  feet. 

A  Room  System  for  Gaseous  Mines. — In  Fig.  17  is  shown 
the  Brown  panel  system  employed  in  thick  seams  of  coal  where 
there  is  considerable  gas  and  the  roof  poor.  At  distances  along 
the  gangway  of  180  feet  are  turned  double  rooms  up  on  the  pitch. 
These  are  24  feet  wide,  separated  by  a  pillar  of  15  feet  in  thick- 
ness, which  latter  is  withdrawn  as  soon  as  the  rooms  have  reached 
the  upper  airway.  Midway  between  each  pair  of  rooms  is  driven 
a  central  heading  12  feet  wide  to  the  full  length  as  a  travelling 
road  and  airway,  as  well  as  a  coal-chute.  From  this,  to  right  and 
left  at  intervals  of  30  feet,  are  turned  similar  headings  carried 
through  the  coal  until  the  double  rooms  are  reached.  The 
coal  is  mined  from  the  rooms  toward  the  central  heading  as 
indicated  in  the  figure,  and  in  such  order  as  to  leave  flanking 
pillars  as  shown.  Small  cars  convey  the  coal  to  the  chute.  A 
heavy  barrier  pillar  separates  each  pair  of  rooms,  and  is  untouched 
until  the  boundary  of  the  property  has  been  reached.  The  sys- 


62 


MANUAL  OF  MINING. 


METHODS   OF  MINING.  6j 

tern  affords  safety  to  the  men,  but  is  expensive  because  of  the 
large  amount  of  narrow  work  performed  before  much  coal  is  pro- 
duced. 

The  County  of  Durham  System  is  a  combination  of  the  panel 
and  the  pillar- and- room  methods.  The  breasts  with  their  pillars 
are  laid  in  groups  of  eight  or  ten  from  the  gangway,  each  section 
or  group  being  mined  separately  and  systematically  toward  the 
boundary.  Between  the  sections  are  left  barrier-blocks  of  coal 
150  feet  thick,  unbroken  except  for  airways.  These  give  com- 
plete isolation  to  the  section  and  also  localize  whatever  move- 
ment may  occur  in  the  falling  of  the  roof.  As  the  coal  in  these 
barrier-pillars  suffers  in  quality  from  the  crush  of  the  roof,  their 
location  is  selected  where  the  coal  is  of  poor  quality.  They  are 
not  touched  until  all  of  the  adjacent  sections  have  been  exhausted. 
This  method  is  safer  than  the  pillar  or  room  in  gaseous  mines, 
and  more  economical  than  the  panel  method. 

Anthracite  Mines. — The  mining  of  anthracite  presents  condi- 
tions that  are  not  usual  in  other  fields.  The  contortions  and 
folds  to  which  the  strata  have  been  subject  during  the  processes 
of  mountain-making  have  developed  so  many  eccentricities  of 
pitch  that  it  is  difficult  to  plan  the  works  a  great  distance  in 
advance.  For  example,  in  a  shaft-mine  in  which  the  workings 
may  have  continued  horizontally  for  some  time,  the  coal  might 
suddenly  rise  or  dip  and  change  the  entire  topography,  and  hence 
necessitate  an  entire  revision  of  the  system.  The  system  gener- 
ally adopted  in  the  anthracite  region  is  the  pillar-and-room  on 
the  double-entry  plan.  Levels  are  driven  each  way  from  the 
shafts  sunk  to  intersect  one  or  more  beds.  The  coal  is  invari- 
ably worked  from  the  lower  toward  the  higher  level,  and  the  pillars 
are  shot  through,  leaving  chain-pillars  as  a  solid  support  for  the 
level  above.  The  width  of  the  rooms  and  of  the  pillars  is  not 
materially  affected  by  the  pitch  of  the  beds,  though,  at  times, 
elaborate  timbering  may  be  necessary  to  safeguard  the  men  and 
prevent  the  coal  from  running  into  the  gangways. 

Rock-chute  Mining.  —  In  the  Wyoming  and  Lacka wanna 
valleys,  where  several  coal-seams  are  worked  simultaneously, 


64  MANUAL  OF  MINING. 

special  care  must  be  exercised  that  the  rooms  in  the  different  seams 
are  laid  off  in  such  manner  as  to  leave  their  pillars  in  the  same 
vertical  line.  When  the  beds  are  thick  and  separated  by  partings 
of  slate  of  but  a  few  feet,  or  the  several  beds  are  thin  and  separated 
by  a  few  yards  of  rock,  they  are  mined  simultaneously,  all  being 
connected  by  a  chute  built  from  the  lower  to  the  upper  bed  for 
delivery  of  the  product  to  the  lower  seam. 

Systems  of  Mining  the  Mammoth  Coal-seam  of  Pennsylvania. — 
The  system  followed  at  Hazelton,  Pennsylvania,  in  mining  the 
Mammoth  bed  at  a  point  where  it  is  40  feet  thick,  with  an  inclina- 
tion at  tunes  of  as  much  as  60°  to  the  horizontal,  is  shown  in 
Fig.  13.  This  is  known  as  a  single-chute  breast  method.  The  plan 
shows  the  gangway  a  heavily  timbered  through  the  coal,  and  a 
chute  b  leading  from  the  battery  platform  to  the  gangway.  The 
battery  c  is  made  of  open  timbers  hitched  into  the  floor  and  roof, 
backed  by  horizontal  timbers.  In  it  is  a  door  of  planks  capable 
of  being  removed  when  desired  for  loading  the  cars.  A  chute 
in  the  middle  of  the  room  receives  the  coal,  and  manways,  dd, 
on  either  side,  made  by  leaning  posts  diagonally  from  the  floor 
to  the  rib  of  coal,  provide  air  and  entry  for  the  men.  These  man- 
ways are  often  termed  "  juggler- ways. "  A  breakthrough,  e,  is 
made  between  the  rooms  for  ventilation,  and  the  main  airway,  /, 
is  driven  over  the  gangway,  as  usual,  near  the  roof.  Every  third 
or  fourth  room  is  connected  with  the  intake  and  return  airway, 
and  in  extreme  cases  every  two  rooms  are  also  connected.  The 
amount  of  coal  which  is  withdrawn  from  the  battery  is  only  suffi- 
cient to  relieve  the  battery  of  the  excessive  volume,  sufficient 
being  allowed  to  remain  for  the  miners  to  stand  upon  while  engaged 
in  work. 

A  modification  of  this  single- chute  breast  method  consists 
in  providing  each  room  with  two  batteries,  the  manways  dd 
serving  as  chutes,  the  space  between  them  being  used  for  refuse. 
This  is  the  plan  employed  when  a  large  amount  of  bony  coal  and 
slate  is  present  in  the  seam. 

Propping  through  a  Weak  Floor  and  Roof. — At  the  Richard 
mine,  where  both  the  floor  and  the  roof  are  weak,  a  variation,  as  in 


METHODS  OF  MINING.  6c; 

r-> 

Fig.  1 8,  has  been  introduced.  The  props  are  placed  6  feet 
apart,  with  their  length  such  as  to  drive  them  through  the  weak 
shale  to  the  solid  rock  beneath  the  floor  and  likewise  above  the 
roof  slate.  By  this  plan  an  ample  support  is  given  while  the; 
pillars  are  being  extracted. 


FIG.  18. — Variation  of  System  for  Pillar-robbing. 

The  Room  System  with  Caving. — In  the  North  Staffordshire; 
mines  is  a  system  similar  to  the  one  mentioned  above,  in  which 
the  gob  follows  downward  with  the  mining  and  requires  no  atten- 
tion. This  has  also  been  employed  in  California,  and  resembles 
closely  the  caving  system  as  practised  in  the  iron-ore  mines  of 
Lake  Superior,  as  described  later  on  in  this  chapter. 

A  Room  System  with  Filling. — At  the  Shamrock  colliery,; 
Westphalia,  is  practised  another  system  of  mining  coal  at  45°,  the: 
bed  having  a  thickness  of  seven  and  one  half  feet.  Here  the  bed  isi: 
divided  into  blocks  2000  feet  long  on  the  strike  and  600  feet  ort 
the  pitch.  Each  block  is  subdivided  into  three  levels  (aa,  Fig:  ig), 
200  feet  apart,  through  which  is  driven  the  central  heading  a  for 


66  MANUAL  OF  MINING. 

ventilation.     On  each  side  of  the  pillars  are  self-acting  inclines, 
or  coal-chutes,  ee. 

This  method  contemplates  the  filling  of  the  excavation  made 
by  the  removal  of  the  coal  by  such  material  as  can  be  obtained  at 
the  surface.  The  central  headings  aa  serve  for  lowering  the 
material  used  for  storage.  At  the  point  c  is  a  small  shaft  in  which 
is  accumulated  the  filling  material  and  such  rock  as  is  obtained 
in  taking  up  the  floors  of  the  haulage-ways.  From  this  point  is 


FlG.  19. — A  Room-and-pillar  System  adopted  in  Westphalia. 

obtained  the  material  which  is  delivered  to  the  rooms  through  the 
central  heading  aa  when  needed. 

The  plan  of  the  mine  provides  three  working  places  on  each 
side  of  the  central  heading,  and  during  operations  three  of  them 
are  being  filled,  while  those  on  the  opposite  side  are  in  the  process 
of  excavation.  This  alternation  of  operations  maintains  regular 
pressure  of  the  roof. 

The  coal  is  removed  by  a  process  of  overhead  stoping  (Fig.  27), 
delivered  to  the  levels  a,  and  thence  by  the  inclines  e  to  the  main 
levels.  On  either  side  of  the  central  heading  is  left  the  stump- 
pillar,  which  is  eventually  withdrawn  when  the  level  is  to  be 


METHODS  OF  MINING.  6/ 

abandoned.     Its  lower  end  is  very  carefully  blocked  to  prevent 
the  leakage  of  air  from  the  ventilating  current. 

Mining  Thick,  Steep  Coal-seams  with  Rooms. — The  method  of 
mining  shown  in  Fig.  20  is  described  in  the  Coal-  and  Metal-Miners" 
Pocket-Book  as  having  originated  in  California.  The  coal-seam  is 
7  feet  thick  and  has  an  inclination  of  60°.  The  gangway  a  is  driven, 
along  the  strike  40  feet  from  the  airway  6;  breakthroughs,  cc,  at 
distances  of  30  feet  connect  the  two,  acting  first  as  airways  and 


FlG.  20. — A  Calif ornian  Variation  of  the  Pillar-and-room  System.    . .     ;Q. 

later  as  mill-holes  or  chutes  down  which  to  lower  the  coal.  The 
door  or  gate  at  the  foot  of  each  chute  permits  the  coal  to  be  loaded 
directly  into  the  cars  on  the  gangway.  Counter  airway-chutes 
dd  are  driven  over  30  feet  apart  at  an  angle  of  35°  for  delivering 
the  coal  to  the  chutes,  cc,  by  gravity.  Numerous  breakthroughs, 
40  feet  apart,  divide  the  coal  and  the  lift  between  each  two  pairs 
of  levels  into  pillars.  When  a  panel  of  five  or  more  chutes  has 
been  driven  the  pillars  are  robbed  on  the  retreat  from  one  cor- 
ner at  the  top.  This  is  possible  only  in  comparatively  strong 
coal  without  partings. 

Square  Work. — Square  work  is  a  modification  of  the  pillar- 
and-room  system,  in  which  very  thick  beds,  not  only  of  coal  but 
also  of  salt  gypsum,  puzzolana,  etc.,  may  be  mined.  Rooms 


$8  MANUAL  OF  MINING. 

•are  opened  from  the  gangway  150  feet  square,  with  pillars  30 
.feet  thick  between  them. 

Two  sets  of  cross-galleries  20  feet  wide  are  driven  through 
'the  rooms,  leaving  pillars  about  25  feet  square  for  the  support 
-of  the  roof.  They  are  driven  as  high  as  the  vein-matter  will 
allow.  If  the  roof  or  floor  is  not  firm,  a  layer  of  mineral  may 
ie  left  for  greater  security.  The  objection  to  the  method  is  in 
the  difficulty  of  obtaining  systematic  ventilation.  Moreover, 
<very  little  of  the  pillars  is  recovered,  and  the  system  involves  great 
risk  to  life,  since  the  roof  is  usually  too  far  for  the  eye  to  detect 
any  threatening  disaster. 

Flushing. — About  15  per  cent  of  the  coal  obtained  from  the 
anthracite  mines  is  of  such  a  nature  that  it  requires  special  treat- 
ment, which  may  not  warrant  the  expense  of  attempting  to 
recover  it. 

Moreover,  in  breaking  and  sizing  the  coal  for  the  market  a 
large  quantity  of  fine  coal  is  produced  and  slate  sorted  which 
were  formerly  dumped  and  accumulated  in  large  piles,  encum- 
bering the  surface  and  increasing  the  pressure  below.  These 
culm  piles  are  now  being  utilized  for  the  filling  of  the  excava- 
tions by  a  process  known  as  flushing.  In  Scranton,  Pa.,  a  stream 
•of  water  is  discharged  against  the  culm  piles,  washing  the  mate- 
rial into  a  pipe-line  and  down  holes  driven  from  the  surface 
to  the  underground  workings.  The  mouth  of  the  rooms  being 
boarded  up  at  the  gangway  in  such  a  manner  as  to  drain  off 
the  water  without  allowing  any  culm  to  escape,  the  room  becomes 
filled  with  wet  culm.  The  culm  packs  quickly  and  solidly  and 
•offers  a  degree  of  solidity  almost  equal  to  that  of  the  untouched 
-coal  in  the  pillar.  The  pillars  are  then  attacked  on  either  side 
with  safety  and  their  contents  almost  totally  recovered.  The 
roof  is  then  allowed  to  close  in  the  space  and  rest  on  the  floor  and 
culm. 

Metal-mining. — The  location  of  the  point  of  attack  of  a  metal- 
jnine  involves  considerations  which  were  discussed  in  Chapter  II. 
The  comparative  merits  of  the  shaft  driven  on  the  wall  side 
<of  the  vein,  and  of  that  intersecting  the  vein,  as  illustrated  in 


METHODS  OF  MINING. 


69 


Figs.  22  and  23,  were  likewise  reviewed.  The  determination 
of  the  system  of  mining  depends  largely  on  the  capital  avail- 
able and  the  value  of  the  deposit.  In  precious-metal  mining 


FlG.  21. — Following  Vein  by  Incline. 

the  mineral  cannot  be  worked  from  level  to  level  as  rapidly 
as  the  levels  are  pushed  forward,  for  it  is  difficult  to  estimate 
in  advance  the  quantity  of  "pay  ore."  Hence  for  economical 


FIG.  22. 


FIG.  23. 


The  Two  Positions  for  a  Shaft  Relative  to  the  Vein. 

work  only  sufficient  ore  is  extracted  to  pay  the  expense  of 
operation  until  the  limit  of  the  property  has  been  reached. 
This  assumes  on  the  part  of  the  operators  a  degree  of  patience 
which  is  not  usually  found  in  speculative  districts.  Under  ordi- 


70  MANUAL  OF  MINING. 

nary  circumstances  it  is  desirable  or  may  be  necessary  to  obtain 
immediate  returns.  In  Fig.  24  is  shown  the  method  pursued 
in  blocking  out  the  mine.  It  represents  a  longitudinal  view  of 
the  vein.  Through  the  vein-rock  extend  streaks  of  ore  of 
high  grade  which  are  neither  uniform  in  dimensions  nor  in  pitch. 
These  may  be  encountered  at  any  place  in  the  vein-matter.  In 
such  cases  the  risers  a  are  driven,  or  winzers  b  are  sunk,  in  such 
of  them  as  promise  profit.  If  the  ore  is  uniform  in  grade  and 
value,  the  mill-holes  are  spaced  uniformly.  These  serve  to 
develop  the  property,  enabling  the  future  to  be  calculated  with 


FIG.  24.— A  Vein  Blocked  Out,  showing  Ore-shoots. 

some  degree  of  certainty.  The  mill-holes  serve  for  ventilation 
as  well  as  ore-chutes.  Risers  are  cheaper  to  open  than  winzes, 
and  are  preferred  in  wet  mines,  though  they  are  at  a  disadvantage 
in  the  hot  mines. 

Systems  of  Mining  Metalliferous  Veins. — The  precious  metals 
occur  usually  in  thin  veins  of  hard  rock,  and  constitute  a  small 
portion  of  the  deposit,  and  the  systems  adopted  for  their  extraction 
are  known  as  underhand  or  overhand  sloping. 

In  overhand  stoping  the  mineral  is  recovered  by  upward 
workings  in  a  series  of  steps,  the  barren  rock  being  stowed  away 
in  the  lower  portion  of  the  lift.  In  underhand  stoping  the  ore 
is  mined  out  in  steps  working  downward,  while  the  waste  rock 
is  disposed  of  on  platforms  built  to  receive  it. 


METHODS  OF  MINING.  71 

For  thick  vertical  deposits  of  iron,  copper,  lead,  zinc,  or 
other  ores,  three  systems  are  in  vogue,  varying  with  the  strength 
of  the  ore.  The  square-set  and  filling  systems  are  used  in 
large  deposits  and  veins  for  removing  the  entire  contents,  whether 
they  be  friable  or  strong.  The  caving  system  is  designed  for  ore 
in  large  masses  friable  enough  to  break  and  fall  when  undermined. 
The  order  of  proceeding  in  all  these  systems  is  upward.  A 
descending  system  is  employed  when  a  heavy  cover  can  be  utilized 
as  a  flexible  roof,  which  is  allowed  to  cave  with  the  downward 
progress  of  the  excavation. 

The  Stoping  Systems  of  Mining. — After  the  vein  has  been 
blocked  out,  then  such  blocks  as  are  deemed  workable  are  mined 
by  one  of  two  methods,  overhand  or  underhand.  In  the  first 
method  the  miner  picks  or  shoots  down  the  ore  in  front  of  and  above 
him,  advancing  the  breast  as  high  as  he  can  reach  and  as  wide  as 
the  vein  and  as  far  as  the  limit  of  the  block.  In  underhand  work 
the  miner  removes  the  mineral  below  him  and  advances  the  breasts 
longitudinally,  with  the  load  taking  them  successively  downward. 
These  two  methods  are  applicable  to  steep  veins  if  the  overhand 
working  is  quickly  adjusted  in  the  thick  flat  beds.  In  the  former 
method  they  are  merely  divided  from  the  falling  rock,  as  gravity 
facilitates  all  operations  of  breaking  down  the  vein-matter;  this 
is  a  much  quicker,  cheaper,  and  less  arduous  method  than  that  of 
underhand,  as  the  miner  is  compelled  to  raise  the  mineral  before 
the  next  lower  can  be  attacked. 

The  Underhand-stoping  System. — This  is  a  method  of  limited 
application.  When  the  vein  has  been  blocked  out  and  the  several 
blocks  on  any  given  level  are  attacked,  they  are  worked  away  in 
horizontal  slices,  beginning  at  the  winzes  and  extending  half  way 
on  either  side  to  the  adjoining  winzes.  The  miner  standing  on  the 
floor  of  the  level  (Fig.  25)  removes  the  vein-matter  for  a  depth 
of  nearly  6  feet,  sorts  out  the  ore,  and  throws  upon  the  platform 
behind  him  the  waste.  When  these  slices  have  been  carried  as 
far  as  designed,  the  next  lower  is  attacked,  beginning  at  the  winzes, 
its  ore  being  raised  to  the  proper  level  and  its  gangue  thrown  upon 
the  platform  behind  the  miner.  The  platforms  are  constructed 


72  MANUAL  OF  MINING. 

AS  rapidly  as  the  miner  advances,  there  being  one  for  each  hori- 
zontal slice.  If  the  entire  contents  of  the  vein  from  wall  to  wall 
.are  composed  of  pay  dirt,  the  timbering  may  be  dispensed  with. 
It  frequently  happens  that  instead  of  a  single  gang  operat- 
ing in  a  given  block  on  one  slice  at  a  time,  as  many  gangs  are  dis- 
posed in  the  work  as  there  are  slices,  the  uppermost  gang  preceding 
the  one  below,  in  each  case  by  an  amount  sufficient  to  secure  for 
him  the  advantage  of  the  firm  footing  on  the  ore.  Thus  the 
mine  progresses,  as  illustrated  in  Fig.  25,  in  which  the  block  resem- 
.bles  a  set  of  steps  downward  from  the  level  above,  whence  the 


FIG.  25. — Underhand  Stopes. 

:name  underhand  stoping.  Occasionally  the  winze  is  sunk  to  the 
lower  level  at  once,  and  each  step  is  advanced  as  rapidly  as  possible, 
the  mineral  from  all  the  slices  or  steps  being  delivered  downward 
to  the  level  below,  instead  of  requiring  to  be  raised. 

This  system  of  underhand  stoping  is  economical  and  has 
many  advantages  as  compared  with  the  overhand  system  to  be 
described  later.  Timber  is  used  only  for  the  support  of  the  frag- 
ments of  rock  and  for  building  platforms.  The  ventilation  is 
good,  and  better  drilling  can  be  conducted  by  the  men  in  under- 
hand than  in  overhand. 

The  iron-ore  deposits  of  the  Eastern  States  are  worked  down- 
ward, the  vein  being  entirely  removed  except  for  the  occasional 
pillars  left  to  support  the  walls.  The  deposits  vary  in  thickness, 
the  average  being  probably  12  feet,  and  in  pitch  between  15° 


METHODS  OF  MINING. 


73 


and  70°.  In  Fig.  26  the  shaft  pillars  ar^  at  aa;  those  at  bb 
support  the  roof;  at  cc,  the  surface,  and  at  gg,  the  track  in  the 
upper  level.  At  dd  is  the  untouched  ore.  The  latter  is  broken 
down  from  the  stopes,  eet  with  air-drills  and  loaded  into  cars 
below.  The  levels  are  60  feet  apart.  Each  stope  can  furnish 


^xxx^x 

FIG.  26. — The  Underhand  Sloping  of  Iron  Ore. 

about  50  tons  in  a  single  shift.  The  hoisting-car  track  follows 
the  foot-wall  except  where  it  is  too  irregular,  when  the  hollows 
are  bridged  over  or  the  projections  cut  away,  as  at  ft,  to  obtain  a 
regular  slope. 

The  Overhand  Stoping.— The  ore  in  this  case  is  mined  from 
below  upwards,  as  shown  in  Figs.  27  and  28.    The  attack  upon 


FIG.  27. — A  Double-wing  Overhead  Stope. 

the  blocks  of  mineral  begins  at  some  point  on  the  upraise,  the 
miners  standing  on  the  caps  of  the  timbering  in  the  gallery,  as  shown 
on  the  extreme  right  or  left  in  Fig.  27.  The  miners  remove  a 


74 


MANUAL  OF  MINING. 


horizontal  slice  for  a  height  of  about  6  feet,  as  far  as  desirable, 
sorting  out  the  ore  and  delivering  it  to  the  mill-hole,  where  the 
chute  discharges  it  into  cars  below.  Meanwhile  the  waste  rock 
is  piled  on  the  gangway  timbers  behind  the  men.  The  next 
upper  s.lice  follows,  and  is  broken  away  in  the  same  manner  from 
the  mill-hole,  the  miners  then  standing  on  the  waste  rock  from  the 
previous  slice.  If  the  vein  can  supply  no  waste  rock,  a  temporary 


""" 

*w   $»  r 

LOA/G/TVD/MAL    SEC  T/ON  Of  <S  TOPE 

FlG.  28.— Longitudinal  Section  of  Stope. 

staging  is  placed  conveniently  for  mining  as  is  shown  near  the 
top  of  the  room  in  Fig.  28.  The  room  in  this  case  presents  an 
appearance  of  inverted  steps,  whence  originates  the  name.  In 
the  double  wing  stopework  advances  simultaneously  from  both 
sides  of  the  mill-hole,  as  in  Fig.  28. 

In  shooting  down  the  ore,  sheet  iron  or  boards  are  laid  on 


METHODS  OF  MINING.  75 

the  floor  to  receive  the  mineral  and  prevent  the  pulverized  and 
friable  ore  from  being  lost  in  among  the  waste. 

Comparison  of  Underhand  with  Overhand  Systems. — The 
comparative  merits  of  these  systems  depend  largely  upon  the 
amount  and  cost  of  timbering.  From  this  point  of  view  the 
overhand  has  the  preference,  as  requiring  little  or  no  timbering 
for  the  reception  of  the  waste.  The  underhand  requires  a  plat- 
form for  each  step  in  each  room.  If  the  vein  is  entirely  salable, 
there  is  then  no  waste,  and  very  little  permanent  timbering 
will  be  needed  for  underhand  stoping.  The  underhand  system 
is  accounted  best  for  valuable  ores  because  all  rock  must  be 
handled,  and  there  is  no  occasion  for  any  being  lost.  On  the 
other  hand,  as  the  ore  is  being  trodden  upon  continually,  it  can- 
not be  used  for  brittle  minerals.  Overhand  work  affords  some 
advantages  in  regard  to  breaking  down  mineral.  In  a  wet  mine 
the  underhand  stoping  presents  objections  unless  the  mill-holes 
and  winzes  have  been  sunk  to  the  lower  level.  The  overhand 
method  is  practicable  in  a  wider  vein  than  is  the  underhand, 
which  is  limited  to  a  width  of  vein  by  the  length  and  size  of  the 
stulls  which  are  available. 

The  Square-set  System.  —  When  the  vein  of  metalliferous 
mineral  is  wider  than  8  feet,  the  underhand-stoping  method  used 
to  introduce  the  timbers  to  reach  from  one  wall  to  the  other  must 
be  exceptionally  long  and  correspondingly  stout,  hence  is  very 
difficult  to  handle  with  the  conveniences  available  underground. 
The  method  then  employed  may  consist. in  the  substitution  of  a  set 
of  frame-timbers  (Fig.  258).  This  method  of  square-set  timber- 
ing is  applicable  to  either  hard  or  soft  ore.  It  is  equally  applicable 
to  flat  as  to  steep  deposits,  and  succeeds  just  as  well  with  soft  ore 
as  with  firm  ore. 

The  vein  is  blocked  out  in  the  usual  manner  into  lifts  of  60 
feet  each  by  gangways  or  levels  running  longitudinally  in  the 
vein,  if  its  contents  are  of  hard  rock,  or  in  the  country  rock  if  soft. 

Along  the  gangways  are  alternate  rooms  and  pillars,  each  of 
the  same  width,  from  20  to  40  feet  long,  and  as  wide  as  the  deposit. 
The  breast  of  each  room  is  divided  into  working-faces  of  about 


76  MANUAL  OF  MINING. 

7  feet  wide.  The  erection  of  the  square  set  of  timbering  begins  as 
soon  as  possible,  props  and  wedges  being  placed  between  the 
timbers  of  the  set  and  the  roof  or  side  rt>ck  to  prevent  fall.  With 
the  advance  of  the  room  additional  timbers  are  laid  in  the  frame 
on  the  floor,  at  the  side,  in  front  and  above ;  thus,  as  the  room  is 
being  excavated  to  its  limit,  the  framing  assumes  a  form  similar  to 
that  in  Fig.  258.  • 

The  square  setting  of  this  type  is  employed  in  mining  cham- 
bers or  deposits  when  the  ore  is  of  a  comparatively  uniform 
nature.  It  is  most  economical,  for  the  steps  can  be  worked  on 
all  four  sides.  When  the  first  sets  are  fully  under  way  for  a  con- 
siderable area,  the  first  floor  is  stoped  out  and  sets  placed  directly 
over  the  sill-sets,  care  being  taken  to  be  at  least  two  sets  behind 
those  on  the  first  floor,  in  order  to  keep  sufficient  ore  on  the  floor 
to  work  with  while  the  sets  are  being  placed  in  position. 

The  operations  of  mining  may  be  confined  to  one  level  only,  or 
they  may  be  prosecuted  simultaneously  from  several  levels  working 
across  the  vein  and  upward  in  each  lift.  The  mining  proceeds  by 
slices,  ascending  to  the  upper  level,  as  in  the  systems  of  sloping,  the 
timbers  of  the  respective  sets  being  placed  in  vertical  tiers.  The 
mining  is  also  conducted  in  an  order  resembling  longwall,  in  which 
the  entire  face  is  attacked  and  the  timbering  placed  at  once,  the 
progress  then  being  in  every  direction  except  downward.  This 
modification  is  favored  where  the  mineral  is  of  low  value  and  very 
soft. 

When  the  rooms  on  a  level  have  been  exhausted  and  the  timbers 
carried  to  the  upper  level,  work  begins  upon  the  intermediate 
pillars  in  the  same  manner  as  practised  in  the  rooms. 

In  Fig.  29  is  illustrated  the  method  of  sloping  with  square  sets 
where  the  work  is  prosecuted  from  the  hanging-wall  side  of  the 
deposit  and  carried  across  it.  This  plan  assumes  the  hanging-wall 
to  be  heavy  and  liable  to  cave  or  the  ore  soft.  The  construction 
of  the  square  sets,  b,  is  then  conducted  as  rapidly  as  sufficient 
space  is  excavated  to  permit  of  it.  The  set  c  is  next  provided  for 
and  erected,  after  which  the  ground  above  them  is  stoped  out  for 
the  high  set  d  and  the  flat  set  g.  The  set  e  is  next  erected,  as  shown 


METHODS  OF  MINING. 


77 


at  /,  and  the  sets  built  up  in  succession  until  the  hanging-wall 
is  reached. 

To  avoid  any  risk  of  movement,  care  must  be  exercised  that 
the  posts  be  perfectly  in  the  line  of  pressures  and  are  securely 


FlG.  29. — The  Square-set  System,  working  from  the  Hanging-wall. 

blocked  by  wedges  at  the  points  of  contact  with  the  hanging-wall. 
Where  there  is  any  pressure  the  sets  are  reinforced  by  braces,  as 
at  h.  Occasionally  ladders  ii  are  carried  up  for  the  men,  and 
chutes  j  arranged  for  the  delivery  of  the  ore  into  the  bin.  All  ore 
is  removed  from  the  vein.  The  support  for  the  roof  depends  upon 
the  timbering  alone  and  upon  its  security.  Care  must  therefore 
be  taken  to  prevent  caves. 

In  Fig.  30  is  an  illustration  of  square- set  timbering  beginning 
at  the  foot-wall.    The  rock  is  first  cut  for  the  sills  a,  which  are 


78  MANUAL  OF  MINING. 

placed  in  position,  and  the  sills  &,  of  extra  lengths  to  secure  a  firm 
base.  Sometimes  they  are  of  especial  lengths,  as  at  d.  Hitches 
are  cut  directly  under  the  posts  of  the  set  above  and  the  short 
posts  c,  the  caps  then  having  different  lengths  to  simplify  the 
framing. 

When    the    deposit    contains    considerable    barren    material 
which  can  be  employed  as  waste  and  thrown  into  the  spaces 


FlG.  30. — The  Square-set  System,  working  from  the  Foot-wait 

between  the  sets  to  accumulate  and  eventually  assist  the  timbers  in 
sustaining  the  pressure,  the  sets  at  the  level  are  reinforced  by 
auxiliary  timbers  g  and  h,  and  also  lagged  to  keep  the  level  open. 
The  mill-hole  m  is  surrounded  by  a  gob  and  serves  as  an  airway. 
At  p  are  props  used  to  temporarily  support  the  ground  and  to 
guard  against  accident  during  the  operations  of  sloping,  and  the 
single  braces  are  introduced  to  prevent  deformation  of  the  frames. 
In  the  Lake  Superior  region  several  of  the  iron-mines  are 
worked  in  this  manner,  the  lower  sets  being  subsequently  filled 
with  rock  or  sand  lowered  into  the  mine  from  above. 


METHODS  OF  MINING. 


79 


Top-slicing  and  Caving. — There  is  a  system  of  mining  which 
has  been  successful  in  wide  deposits  where  the  cover  can  be  induced 
to  break  and  follow  the  excavation  from  a  higher  level  .to  a  lower 
one.  The  mineral  is  removed  from  the  top  of  the  deposit  in  horizon- 
tal slices  8  feet  wide  and  6  feet  high,  made  by  a  series  of  horizontal 
drivages  timbered  with  the  three-stick  sets  (Chapter  III,  Part  II) 
as  the  work  progresses.  On  one  side  is  the  solid  ore,  on  the  other 
is  refuse  of  previous  slices  which  had  caved  when  the  timbers 
were  removed. 

The  shaft  is  sunk  in  the  country  rock,  and  from  it,  at  levels  of 
60  feet  each,  are  driven  cross-cuts  to  the  deposits.  The  first  cross- 
cut is  extended  from  one  wall  to  the  other,  and  from  the  centre  of 
it  are  driven  levels  at  right  angles  to  the  outer  boundary  lines,  at 
intervals  of  100  feet;  on  this  level,  risers  are  carried  up  nearly  to 
the  cover,  from  which  are  run  cross-drifts  until  the  side  walls  of  the 
deposits  have  been  reached.  The  horizontal  slice  is  then  removed 
at  the  top  parallel  to  the  wall  and  the  space  timbered.  Its  length 
on  either  side  of  the  cross-cut  does  not  exceed  50  feet. 

In  Fig.  31  is  shown  this  system,  in  which  a  is  the  main  level, 


FIG.  31. — A  System  of  Top-slicing  the  Ore. 

b  the  riser,  d  the  cross-cut  on  top  of  the  ore,  and  e  the  cave  fol- 
lowing the  slicing.     A  riser  is  made  6  or  8  feet  and  divided  into 
two  compartments,  one  as  a  chute  and  the  other  for  a  ladderway. 
No  effort  is  made  to  support  the  roof,  and  the  cover  is  there- 


8o  MANUAL  OF  MINIXG. 

fore  allowed  to  cave,  the  caving  taking  place  as  the  men  retreat  from 
the  end  of  the  walls  toward  the  risers,  and  from  the  ends  toward 
the  cross-cuts,  and  from  walls  toward  the  risers. 

When  the  excavation  has  been  carried  50  feet  from  the  cross- 
cut, the  timbers  are  withdrawn,  or  cut  or  blown  down,  as  the  case 
may  be,  in  order  to  permit  the  roof  to  cover  as  soon  as  the  work  is 
completed.  Prior  to  removing  the  timbers,  however,  there  are 
placed  on  the  floor,  3  or  4  long,  8- inch  poles  along  the  passages, 
with  split  lagging  or  sawmill  slabs  across  them.  These  support 
the  roof  when  the  ground  caves,  and  enable  the  timber-sets  of 
succeeding  slices  below  to  be  placed  without  great  difficulty. 

In  very  treacherous  vein-matter  or  under  a  bad  roof,  each 
cross-cut  has  risers  at  stated  intervals  that  take  the  place  of  the 
level  for  removing  the  ore  to  the  main  level.  The  remainder  of 
the  work  does  not  differ  materially  from  that  described. 

Subdrifting  and  Caving  System. — This  system  originated  at 
the  Brotherton  mine  in  Lake  Superior,  and  has  been  adopted 
with  excellent  results  in  many  wide  soft-ore  mines.  Cross-cuts  are 
driven  from  the  hanging-  to  the  foot- wall,  where  they  connect  with 
longitudinal  levels.  At  intervals  of  50  feet  risers  a  (Fig.  32)  are 
driven  50  feet  upward  to  the  level  above.  The  first  six  feet  are 
cribbed  and  a  couple  of  driven  sets  put  on  over  them.  Subdrifts 
c  are  then  driven  each  way  to  meet  the  corresponding  drifts  from 
the  adjacent  risers.  The  riser  meanwhile  is  continued  another 
6  feet  upward,  when  timbers  for  a  second  subdrift  d  and  above 
them  for  the  third  subdrift  e  are  laid,  until  finally  the  riser  reaches 
the  level  above  at  a  distance  of  50  feet.  This  divides  the  space 
between  the  two  levels  into  three  subdrifts,  each  having  a  top  or 
back  of  ore  above  them  6  feet  thick.  The  ore  from  subdrifts  is 
delivered  through  a  riser,  and  shipped  to  the  cars  on  the  lower  level. 

Following  the  completion  of  the  subdrifts  the  capping  of  ore 
above  the  level  g  is  next  removed  by  dividing  the  ore  into  blocks 
of  10  feet  square,  beginning  at  the  end  of  the  level  g.  This  weakens 
the  pillars,  which  crush  and  permit  the  removal  of  the  ore.  That 
above  the  subdrift  e  is  caved  in  a  similar  manner,  followed  by  d 
and  c. 


METHODS  OF  MINING. 


Si 


Occasionally  it  is  necessary  to  support  temporarily  the  timbers 
above  the  ore,  as  at  h,  by  long  timbers,  but  in  all  cases  all  the  close 
timbering  is  employed  to  control  the  pressure  and  prevent  accident. 
In  this  method  all  timber  can  be  lowered  from  the  levels  above 
and  handled  without  interference  with  haulage.  Ventilation  is 
good,  and  simple  means  of  escape  are  provided  in  case  of  accident. 


FIG.  32.— Subdrifting  and  Caving. 

Slicing  and  Filling  System. — This  system  is  practised  in  wide 
deposits,  and  consists  in  taking  two  slices  of  mineral  in  successive 
squares  from  a  lower  to  a  higher  level  and  filling  up  the  excavation 
with  rock  or  other  material  quarried  for  the  purpose.  In  this 
method  the  lifts  are  60  feet  each  and  the  slices  of  attack  6  to  7 
feet  high,  taken  horizontally  across  the  vein  for  as  broad  a  face  of 
attack  as  may  be  decided  upon.  Two  main  levels  c  and  d  (Fig.  33) 
extend  to  the  end  of  the  district  to  be  thus  mined,  being  at  intervals 
connected  "by  risers  e.  The  first  slice  is  removed  at  level  c,  and 


82  MANUAL  OF  MINING. 

waste  rock  from  above  delivered  through  the  opening  e,  which  is 
then  spread  to  fill  the  excavation  completely.  The  temporary 
level  /  is  made  on  the  top  of  the  setting  g.  In  each  successive  slice 
chutes  are  constructed  for  the  purpose  of  letting  down  mineral  into 
the  lower  level,  from  which  it  may  be  trammed  to  the  shaft.  This 
process  is  repeated  until  all  of  the  mineral  has  been  worked  out  in 


FIG.  33.— A  Slicing  and  Filling  System. 

slices  represented  by  dotted  lines  and  the  waste  rock  distributed 
to  fill  the  spaces. 

In  Fig.  34  is  shown  another  method  in  which,  if  the  exploration 
of  the  deposit  is  determined,  this  form  of  slicing-shaft,  number 
8,  may  be  located  permanently,  and  the  rock-shaft,  number  7, 
sunk.  The  drift  A  on  the  hanging-wall  side  of  the  deposit  is 
connected  by  cross-cuts  with  the  level  C.  Cross-cuts  are  also 
driven  through  the  deposits  for  the  drift  A  to  the  foot-\vall  and 
winzes  sunk  between  the  levels.  The  deposit  having  thus  been 
blocked  out,  in  each  lift  of  60  feet  high  there  are  three  methods 
of  removing  the  ore,  according  to  the  extent  of  the  face,  which 
may  prove  to  be  self-supporting. 

First  Method.— From  the  drift  A,  slices  DD  are  taken  6  feet 
wide  on  either  side  of  the  timbering,  cross-cut  B,  whose  space 
is  being  filled  with  rock  dumped  into  A  through  a  winze  from  the 


METHODS  OF  MINING.  83 

level  above.  The  next  two  slices,  EE,  are  taken  while  the  excava- 
tions DD  are  being  filled  and  the  slices  FF  are  mined,  while  E  is 
being  filled  with  rock.  The  filling  material  is  lowered  from  the 
surface  through  rock-shaft  number  7  and  is  hauled  from  the  shaft 
to  the  winzes  above. 

This  process  is  continued  until  a  layer  6  feet  thick  has  been 
removed  over  the  entire  deposit.    The  next  slice  above  is  operated 


FIG  34.— The  Filling  System,  Plan  of  One  Level. 

in  the  same  manner,  with  the  exception  of  a  new  drift  driven 
above  A ,  as  in  timbering  as  at  B.  The  miners  in  removing  the 
second  slice  stand  on  the  waste  while  at  work,  and  behind  them 
is  spread  the  material  lowered  for  the  purpose.  The  fillers  follow 
the  miners  very  closely.  Mill-holes  are  timbered  upward  as 
mining  progresses,  the  filling  being  packed  about  them. 

Second  Method. — When  the  vein-matter  is  dry  and  stands 
firmly  it  admits  of  a  wider  vein  than  6  feet;  the  attack  may  be 
made  from  the  foot  of  the  wall  side  for  a  width  of  half  the  distance 
between  the  cross-cuts  BB.  The  ore  is  then  removed  through  the 
cross-cut  levels  while  the  filling  is  being  spread  as  above  described. 
The  slices  are  driven  in  a  direction  parallel  to  the  strike  and  across 
it,  as  in  the  first  method.  It  may  be  well  to  mention  that  similar 


84  OF  MINING. 

operations  are  being  conducted  not  only  at  other  points  on  the 
same  level,  but  also  in  a  similar  manner  at  other  levels  above  or 
below. 

Third  Method. — This  is  designed  for  veins  not  over  20  feet 
wide.  Here  the  face  of  the  attack  is  the  two  sides  of  the  pillars 
cut  out  by  cross-cuts  BB.  The  miners,  starting  from  the  cross-cuts, 
work  toward  one  another,  parallel  with  the  strike,  until  they  meet, 
their  first  cut  being  made  at  the  walls,  the  filling  progressing 
behind  them  as  before. 

This  system  of  filling  is  considered  a  cheaper  and  safer  system 
than  either  caving  or  square  set  for  the  extraction  of  mineral, 
whether  soft  or  hard,  in  large  deposits,  particularly  if  a  neigh- 
boring quany  can  supply  the  waste  rock.  It  is  of  the  widest  range 
of  application.  Very  little  timber  is  needed,  and,  with  the  excep- 
tion of  the  first  cross-cut  B  made  in  the  solid  ore,  the  mining  is 
cheaply  done,  because  the  width  of  face  is  large  and  the  miners 
have  two  free  faces  to  take  advantage  of  while  blasting.  Any 
material  whatsoever  will  serve  for  filling,  but  the  best  is  that 
mixed  with  clay,  as  it  packs  well.  It  should  be  in  pieces  as  large 
as  may  be  conveniently  handled,  for  the  "smalls"  settle  too  much. 
In  a  few  weeks'  time  the  entire  mass  of  filling  compacts  so  that  no 
difficulty  is  experienced  when  the  stowing  of  the  uppnr  level  is 
reached. 

In  the  caving  system,  extreme  caution  is  necessary  to  prevent 
sudden  falls  of  roof,  and  the  adoption  of  the  method  is  only  justi- 
fied where  there  is  an  absence  of  filling  material  and  a  scarcity  of 
timber.  But  if  the  vein-matter  is  comparatively  firm,  the  method 
with  filling  presents  greater  security,  and  may  be  cheaper.  In 
such  event  the  choice  of  method  depends  upon  the  relative  expense 
of  procuring  quarried  rock  and  framed  timber. 


METHODS  OF  MINING.  8$ 


REFERENCES. 

Mining  Thin  Seams  at  Escatpelle,  Coll.  Guard.,  LXXII,  365;  France 
and  Belgium,  Coll.  Guard.,  LXXII,  Serial;  North  Staffordshire  Fed.  Inst. 
M.  E.,  VIII;  Mining  Thick  Seams  in  Belgium,  Coll.  Guard.,  June  1896, 
1202;  Mining  Methods,  Pratt  Coal,  A.  I.  M.  E.,  XIX,  299;  Connellsville, 
XIII,  330;  Report  A.  C.  Second  Geol.  Survey  of  Penn.  Anthracite;  Report 
Second  Geol.  Survey  of  Penn.  Bituminous,  Annual  Report,  1886;  Mining 
Methods  in  the  Connellsville  Coke  Region,  F.  C.  Keighley,  Eng.  Mag.,  Oct. 
1900;  Mining  Methods  and  Safety  Measures  in  the  Wilczek  Mines  in  Polish 
Ostrawa,  Josef  Mauerhofer,  Oesterr.  Zeitschr.  f.  Berg  u.  Huttenwesen,  May 
16-23,  1903. 

A  New  Method  of  Coal  Mining,  Genie  Civil,  Dec.  16,  1889;  Coal  Mining 
in  Natal,  C.  J.  Gray,  Coll.  Guard.,  Nov.  6,  1903;  Coal  Mining  in  New  South 
Wales,  Coll.  Guard.,  Oct.  23,  1903;  Coal  Production  and  Consumption  of 
the  Principal  Countries  of  the  World,  Coll.  Guard.,  Oct.  30,  1903;  Mining 
Coal  at  the  Face,  J.  T.  Beard,  Mines  &  Min.  Jour.,  April  18,  1903. 

Pillar  and  Stall  Mine  Insp.,  8th  Kans.,  21,  13;  With  Longwall,  Coll. 
Eng.,  Feb.  1897,  317;  Thick  Seam  Anthracite  in  Colo.,  Coll.  Eng.,  April 
1897;  Improvements  Suggested  by  Mine  Insp.,  Coll.,  May,  438;  Boundary 
Pillars  Illust,  Mine  Insp.,  Pa.,  1886;  Pocket,  Shaft  Pillars,  Coll.  Guard., 
June  18, 1897,  1185;  Coal  Pillars,  M.  &  S.  Press,  1893;  Pillars  and  Waste  in 
Coal,  Editorial  and  Historical,  Coll.  Guard.,  Dec.  18, 1896,  1153;  Shaft  Pillars, 
Coll.  Eng.,  July  1897;  Coal  Irruptions  into  Breasts,  Coll.  Manager,  Dec.  18, 
1896,  634;  Coll.  Guard.,  Dec.  1896,  in;  Mine  Creeps,  111.  Min.  Inst.,  I, 
239;  The  Working  of  Seams  and  Faults  in  Coal  Measures,  H.  Fleck,  Glauck- 
auf,  Jan.  3,  1903;  Mine  Development  Methods,  B.  J.  Forrest,  Can.-Min. 
Rev.,  June  30,  1903;  Surface  Subsidences  Caused  by  Mine  Workings,  Coll. 
Guard.,  Sept.  1897,  569;  Suggestions  for  Improvements  of  Methods,  Mine 
Insp.,  Pa.,  in  Anthra.  Fields,  Coll.  Eng.,  May  1897,  438;  Causes  that  Prevent 
Improvements,  Coll.  Eng.,  May  1897,  462. 

Mining  Soft-ore  Bodies,  A.  I.M.  E.,  Vol.  XVI,  862;  L.  Sup.  M.  Inst.,  I., 
13;  Mining  Method  on  the  Mesabi  Range,  C.  E.  Bailey,  A.  I.  M.  E.,  Vol. 
XXVII,  529;  Iron  Mining,  Menominee  Iron  Mines,  A.  I.  M.  E.,  XVI,  891; 
A.  I.  M.  E.,  XVII,  107;  The  Iron-ore  Mines  of  Biscay,  Bennett  H.  Brough, 
Cassier's  Mag.,  April  1903;  New  Jersey  Iron  Mines,  S.  of  M.  Quart.,  Jan. 
1885,  no. 

Tamarack  Mine,  Cassier's  Mag.,  Jan.  1897,  215;  Osceola  Mines,  Coll. 
Eng.,  1895,  217;  Butte,  Montana,  Mineral  Industry,  III,  175. 

Mining  Methods,  Mineral  Industry,  II,  379;  Thick  Deposit  of  Ore, 
S.  of  M.  Quart.,  Jan.  1883,  100;  Zinc  and  Lead  Mines,  Iowa  Geol.  Survey, 


86  MANUAL  OF  MINING. 

VI,  1897;  Zinc  and  Lead  Mines,  Missouri,  Geol.  Survey,  VI  and  VII,  Silver 
Mining  in  South  America,  Cassier's  Mag.,  V,  435;  Methods  of  Mining  and 
Timbering  in  Large  Ore  Bodies  in  British  Columbia  and  Michigan,  Norman 
W.  Parlee,  Can.  Soc.  of  Civ.  Engrs.,  Adv.  Proof,  Feb.  25,  1904;  Mine  Tim- 
bering by  the  Square  Set  System  at  Rossland,  B.  C.,  Bernard  McDonald, 
Can.  Min.  Rev.,  Sept.  30,  1902. 

Coal  Mining  in  New  South  Wales,  Harrison  F.  Bullman,  Coll.  Guard., 
Feb.  23,  1900;  Earth  Pressure  in  Deep  Mines,  Coll.  Guard.,  Vol.  LXXXIV, 
23;  Flushing  Goaves,  ColL  Guard.,  Vol.  LXXXIV,  1274. 


CHAPTER  IV. 

POWER   GENERATION. 

The  Power  Plant. — The  engine  plant  consists  of  boiler  and 
engine,  occasionally  supplemented,  in  favored  localities,  by  water- 
power  electric  plants.  The  plant  must  be  installed  as  compactly 
as  possible  and  of  ample  size.  Common  prudence  suggests  that 
the  machinery  be  in  slight  excess  of  the  immediate  wants.  This 
is  more  important,  the  lower  the  value  of  the  mineral  output. 
While  the  latest  and  best  machinery  will  in  the  long  run  prove  to 
be  more  economical,  the  question  of  finance  and  the  amount  of 
capital  available  is  not  to  be  ignored,  and  the  engineer  may  be 
compelled  to  content  himself  with  expedients  until  the  mine  has 
been  fully  developed. 

The  mechanical  considerations  will  call  for  the  following 
machinery:  In  a  slope  mine,  stationary  engines  or  locomotives  for 
haulage  ;  in  a  shaft  mine,  an  additional  hoister;  cutting  machinery 
for  coal,  or  drills  for  the  rock  or  ore;  fan  for  ventilation;  pumps 
for  the  accumulating  waters;  machinery  in  tipple,  breaker,  or 
mill  for  crushing,  screening,  and  jigging  the  mineral;  pumps  in 
the  mill;  electric  lights  for  the  mill  and  a  portion  of  the  mine. 
Either  steam,  electricity,  or  compressed  air  may  be  the  motor 
agent  in  any  of  these  cases;  their  comparative  advantages  and 
economy  are  discussed  in  Chapter  VI. 

The  fan,  the  steam-locomotives,  and  such  oil-engines  as  may  be 
employed  are  alone  independent  of  the  central  plant.  Other 
machinery  receives  its  power  from  the  central  source.  Except 
the  self-contained  oil-engine  and  the  rare  hydraulic  engine,  all 
other  motors  depend  for  their  primal  power  on  the  steam  gen 
erator. 

87 


S8  MANUAL  OF  MINING. 

The  Boiler  Rating. — The  boiler  is  essentially  a  device  for 
converting  water  at  low  temperature  into  steam  at  a  high  tempera- 
ture, the  measure  of  its  capacity  being  the  quantity  of  water  which 
it  can  evaporate  in  a  given  time.  It  must  have  sufficient  surface 
exposed  to  fire  to  accomplish  this  end,  hence  the  boiler  is  rated 
on  its  evaporative  power  and  the  area  of  its  heating  surface. 
The  commercial  requirements,  demanding  some  definition  of 
boiler  unit  commensurate  with  that  of  engines  which  it  must 
supply,  have  accorded  an  evaporation  capacity  of  34.5  Ibs.  of 
water  per  hour  from  and  to  212°  as  equivalent  to  a  horse-power. 
The  standard  horse-power  corresponds  to  30  Ibs.  (3.55  gallons) 
of  feed  water  evaporated  from  100°  F.  to  steam  at  70  Ibs.  gauge 
pressure  per  square  inch.  This  requires  an  absorption  by  the 
water  through  the  boiler  shell  of  33,305  British  thermal  units 
per  hour,  practically  equivalent  to  the  centennial  unit  just  men- 
tioned. The  general  term  of  unit  horse-power  thus  retained, 
though  a  deceptive  one  and  unscientific,  is  convenient. 

In  calculating  the  size  of  boiler  required  for  a  certain  purpose 
the  amount  of  steam  consumed  per  hour  in  the  aggregate  by  the 
various  engines  may  be  known  from  table,  page  118,  from  the 
given  conditions  of  the  operation.  The  amount  of  heat,  measured 
in  B.T.U.,  necessary  to  develop  the  given  quantity  of  steam  may 
be  obtained  from  the  steam  tables.  The  number  of  boiler  horse- 
power to  give  the  required  service  is  found  by  dividing  the  total 
B.T.U.,  necessary  for  evaporating  the  steam  required  per  hour 
by  33>305  B.T.U. 

This  power  may  be  subdivided  in  any  manner  if  one  boiler  is 
added  in  excess  of  the  average  demand.  It  would  be  more  ad- 
vantageous to  install  the  required  heating  surface  for  average  load, 
remembering  that  any  unit  may  easily  be  forced  33^  per  cent  with- 
out sacrificing  efficiency.  Then  at  times  of  emergency  one  unit  may 
be  repaired  while  the  remaining  units  are  overloaded,  yet  within 
their  capacity.  Thus,  if  800  pounds  of  steam  are  required  per  hour 
as  a  maximum  and  but  3000  pounds  throughout  the  day,  three 
units  may  be  installed.  If  12,000  pounds  of  steam  are  in  demand, 
four  units  may  be  installed,  giving  opportunity  for  one  boiler  to  be 


POWER  GENERATION.  89 

repaired  while  the  remainder  are  forced.  Thus  the  plant  is  not 
too  large. 

Practice  has  shown  that  to  enable  the  absorption  of  33,305 
B.T.U.  to  be  effected  through  the  shell  there  must  be  a  heating 
surface  of  14  sq.  ft.  Hence  14  sq.  ft.  of  surface  exposed  to  the 
flames  under  a  fire-tubular  boiler  constitute  a  boiler  horse-power 
or  12  sq.  ft.  of  exposed  area  in  the  water-tubular  boiler. 

The  grate  area  must  be  sufficient  to  provide  space  for  the 
coal  and  be  at  least  -gV  °f  the  total  heating  surface.  In  other 
words,  each  boiler  horse-power  must  have  from  0.24  to  0.28  sq. 
ft.  of  grate  area.  If  the  grate  is  too  large,  the  temperature  of 
combustion  is  too  low.  High  temperature  in  the  furnace  is  the 
first  condition  of  economy,  the  second  being  that  the  heat  pro- 
duced shall  be  absorbed  as  completely  as  possible  by  plenty  of 
clean  heating  surface.  A  grate  surface  too  small  causes  clinker- 
ing  of  the  coal.  A  low-grade  fuel  requires  a  large  grate.  As  a 
limited  quantity  can  be  burned  on  a  square  foot  of  grate,  the  max- 
imum evaporative  capacity  of  the  boiler  is  limited  by  its  grate 
area. 

The  Standard  of  Boiler  Comparison. — The  equivalent  evapora- 
tion of  a  boiler  is  that  amount  which  would  have  been  evaporated 
had  the  steam  been  produced  at  atmospheric  pressure  (=212°). 
It  may  be  ascertained  by  dividing  the  B.T.U.  necessary  to  boil 
the  feed-water  into  steam  by  966,  the  latent  heat  of  steam  at 
14.7  Ibs.  pressure.  Thus,  feed-water  at  i6o°F.,  converted  into 
steam  at  100  Ibs.  gauge  pressure,  requires  1057  B.T.U.  The 
factor  of  evaporation  is  then  1057-^-966  =  1.094.  The  equivalent 
evaporation  is  1.094  times  the  actual  evaporation. 

The  rate  of  evaporation  is  about  8  to  19  Ibs.  of  feed- water 
per  pound  of  coal. 

The  Feed-water. — Care  must  be  taken  to  secure  ample  supply 
of  pure  water  for  the  boiler.  Acids  in  the  water  corrode  the 
boiler  and  reduce  its  efficiency  and  life.  Salts  in  suspension 
or  in  solution  give  trouble  in  depositing  within  the  boiler  and 
forming  a  compact  layer  which  is  both  injurious  and  wasteful 
of  heat.  A  loo-horse-power  boiler  evaporating  3000  pounds 


MANUAL  OF   MIXIXG. 


"\ 


POWER  GENERATION. 


of  moderately  pure  water  per  hour  would  in  a  year's  time  receive 
248  Ibs.  of  solid  matter,  which  would  cover  the  surface  £  inch 
thick  unless  removed  before  being  consolidated.  Such  a  crust  is 


inevitably  formed  in  all  boilers,  and,  being  a  poor  conductor  of 
heat,  prevents  heat  from  reaching  the  water.  A  scale  of  TV  inch 
thick  on  the  tubes  causes  a  loss  of  13  per  cent  of  the  fuel  heat. 


92  MANUAL  OF  MINING. 

Moreover,  the  boiler-shell  becomes  excessively  heated  for  the 
same  steaming  effects  and  tends  to  blister  and  warp.  The  scale 
may  contract,  and  cracking,  will  suddenly  expose  the  heated 
plate  to  the  water;  the  generation  of  steam  may  follow  and  an 
explosion  ensue. 

Boiler  Scale. — These  two  wasteful  and  dangerous  sources 
require  that  care  be  exercised  in  the  selection  of  the  water-supply 
for  the  boiler  plant.  Mine  waters  are  usually  acid  and  should 
be  neutralized  by  ammonia  or  precipitated  by  chloride  of  barium 
in  the  tank  prior  to  being  delivered  to  the  boiler.  Other  sources 
of  supply,  except  in  the  mountainous  regions  of  the  Rockies  and 
Sierras,  contain  organic  or  mineral  substances  which  should  be 
rendered  innocuous  in  the  boiler  or,  preferably,  removed  before 
introduction  by  feeding  the  water-supply  through  the  econo- 
mizer or  heater,  Fig.  36,  prior  to  its  delivery  to  the  boiler.  A 
considerable  deposition  will  ensue  from  mere  increase  in  the 
temperature  of  the  water;  the  sediment  there  formed  can  readily 
be  removed.  Such  material  as  still  remains  in  suspension  in 
boiler  must  be  rendered  innocuous  by  an  addition  of  kerosene, 
which  would  cover  the  grains  with  a  film  and  prevent  the  sedi- 
ment compacting  too  densely  on  the  tubes  and  shell.  Washing- 
soda  will  dissolve  the  scale  and  change  its  nature.  The  amount 
of  kerosene  to  be  fed  should  not  exceed  2  quarts  per  week  per  100 
horse-power,  and  of  soda  one  pound  per  week.  An  excessive  use 
of  these  remedies  would  cause  "foaming"  and  deliver  wet  steam 
into  the  cylinder.  The  best  remedy  for  the  formation  of  scale  is 
to  blow  off  and  to  clean  the  boiler  regularly  every  week. 

When  the  condenser  is  used  in  a  steam  plant  its  supply  is 
returned  to  the  boiler,  and  if  lubricating-oil  has  been  used  exces- 
sively in  the  engine  a  very  injurious  mixture  is  introduced  into 
the  boiler.  The  oil  should  be  absorbed  from  the  steam  before 
it  reaches  the  condenser,  or  removed  from  the  water  leaving  the 
condenser  before  delivery  to  the  boiler.  It  is  caught  on  screens 
or  separated  by  a  centrifugal  machine. 

It  is  a  question  of  no  small  matter  for  the  engineer  whose 
water-supply  is  chemically  bad  to  determine  whether  to  incur 


POWER  GENERATION.  93 

the  expense  of  a  special  distillation  plant,  or  to  employ  the  con- 
densed exhaust-steam.  Some  engineers  have  solved  the  problem 
by  using  the  very  smallest  amount  of  lubricating-oil  possible 
in  the  pumps  and  engines,  thus  saving  the  boiler  at  the  expense 
of  wear  upon  the  pistons.  The  author  favors  the  plan  of  purify- 
ing by  chemical  or  thermal  process  such  water  as  requires  it  in 
a  feed-water  heater,  or  an  economizer,  before  delivery  to  the 
boiler.  It  may  even  be  desirable  to  sink  an  Artesian  well  for 
water-supply  if  not  otherwise  attainable,  though  in  that  event 
it  should  not  be  drawn  from  a  limestone  formation. 

The  Feed-water  Heater. — The  feed-water  heater  is  designed 
to  utilize  the  heat  in  the  exhaust- steam  or  the  flue- gases  by  bring- 
ing the  feed-water  supply  in  close  contact  with  them.  It  is 
known  as  an  economizer,  because  only  waste  heat  is  employed 
for  the  purpose  The  simplest  form  of  heater  has  a  tank  of 
matched  dovetailed  boards  in  which  is  placed  the  coil  of  pipe, 
one  end  open,  blowing  into  the  atmosphere,  and  the  other  end  con- 
nected with  the  exhaust  from  the  engine.  The  heat  in  the  steam 
is  communicated  to  water  and  its  temperature  raised,  while  at 
the  same  time  the  carbonic  acid  or  carbonates  in  the  water  are 
released  and  the  salts  deposited.  The  sediment  can  readily  be 
removed,  if  a  hand-plate  be  provided  near  the  bottom. 

A  duplicate  tank  to  this  may  then  receive  the  exhaust- steam 
during  the  process  of  cleaning.  Hay  is  often  placed  in  the  water- 
supply  tanks,  filtration  through  which  would  remove  some  of 
the  muddiness. 

Feed-water  heaters  can  also  be  procured  of  the  closed  type, 
consisting  of  a  steel  cylinder  filled  with  tubes  through  which  the 
steam  flows,  while  the  water  circulates  around  them  or  the  reverse. 
The  benefit  derived  from  the  use  of  the  economizers  increases 
with  the  wastefulness  of  the  boiler,  and  the  need  of  puri- 
fication warrants  their  introduction  when  the  water-supply  is 
chemically  bad.  The  life  of  a  well-built  economizer  is  from 
fifteen  to  twenty  years.  The  average  plant  will  save  at  least 
10  per  cent  of  the  fuel  value.  The  gross  return  on  the  invest- 
ment of  the  economizer  is  48  per  cent  annually  with  coal  at  $5 


94  MANUAL  OF  MINING. 

per  ton,  and  19  per  cent  with  coal  at  $2  per  ton,  assuming  twenty 
hours  of  service  out  of  the  twenty-four.  In  addition  to  the  sav- 
ing of  heat  and  the  perfecting  of  the  waters,  there  is  another 
advantage,  that  the  boiler  receives  water  at  a  high  temperature 
and  is  saved  from  the  straining  effects  which  result  from  cold 
water.  A  small  pump,  or  an  injector,  delivers  the  supply  to  the 
boiler.  The  latter  is  cheaper,  equally  reliable,  and  easily  handled. 

Types  of  Boilers. — It  is  a  matter  of  considerable  difficulty  to 
select  the  best  boiler,  because  of  the  great  variety  now  on  the 
market.  The  three  types,  have  each  their  advocates.  The 
Cornish  type  of  boiler,  with  the  fireplace  in  two  large  tubes, 
has  a  high  efficiency,  and  is  much  liked  where  its  length  is 
no  objection.  The  multitubular  boiler,  with  the  fireplace  below 
the  front  end  and  fitted  with  numerous  tubes  through  which  the 
fire  returns  on  its  way  to  the  chimney,  is  a  most  common  type. 
The  latest  type  is  the  water-tubular  boiler  (Fig.  36),  in  which 
the  fire  and  the  gases  circulate  around  the  tubes;  in  these  the 
water  flows  and  is  converted  to  steam.  These  three  types  have 
their  separate  fields  of  usefulness  and  find  wide  acceptance  in 
engineering  circles.  Their  records  can  easily  be  had,  so  they 
must  be  selected  according  to  local  requirements.  The  relative 
efficiency  of  the  three  types  is  i,  1.6,  and  2;  their  relative  initial 
cost  is  very  nearly  in  the  same  proportion. 

The  Cornish  boiler  utilizes  50  per  cent  of  the  fuel  value  in 
the  production  of  steam;  the  fire- tubular  boiler  absorbs  60  to 
65  per  cent  of  the  heat  generated  in  the  fireplace;  and  the  water- 
tubular  boiler  absorbs  fully  80  per  cent  of  the  fuel  value.  The 
comparative  durability  of  the  three  types  is  about  the  same,  with 
an  advantage  in  favor  of  the  more  expensive  water-tubular 
boilers.  They  all  are  good  steamers,  though  the  last-named 
can  be  most  readily  forced  when  the  emergency  arises. 

Water-tubular  Boilers. — The  steam  pressure  to  which  these 
types  may  be  worked  is  respectively  60  Ibs.,  100  Ibs.,  and  200 
Ibs.  pressure  per  square  inch,  as  a  maximum  for  ordinary  condi- 
tions. Inasmuch  as  the  simplest  method  of  increasing  the  power 
of  a  steam-engine  consists  in  raising  the  pressure  of  the  boiler 


POWER  GENERATION.  95 

steam,  it  follows  that  when  such  emergency  arises  the  water- 
tubular  boiler  will  best  fit  the  case.  These  boilers  are  sectional 
and  can  be  easily  transported.  They  are  safer  than  the  ordinary 
type  even  at  higher  steam  pressure,  because  the  excessive  pres- 
sure beyond  the  danger  line  will  overstrain  but  one  tube  instead 
of  the  entire  boiler  of  the  other  types.  They  are  quickly  cleaned, 
easily  repaired,  and  give  drier  steam  and  with  a  smaller  amount 
of  fuel  than  the  other  types. 

Heat  Losses. — All  the  heat  developed  by  coal  must  be  al> 
sorbed  in  steam  production.  There  are,  however,  three  sources 
of  loss — that  due  to  heat  carried  by  the  flue- gases  escaping  up 
the  chimney,  the  radiation  from  the  walls  and  shell  of  the  boiler, 
and  the  heat  remaining  in  the  ashes. 

The  waste  in  the  ashes  depends  on  the  conditions  in  firing, 
etc.  Their  weight  may  be  20  per  cent  of  the  total  fuel  burned, 
and  in  them  is  also  some  unburned  combustible.  The  unpro- 
tected boiler  surface  radiates,  per  square  foot,  650  B.T.U.  per 
hour.  Thus  the  loss  from  a  100- horse-power  boiler  and  its  wall 
is  equivalent  to  the  heat  of  100  tons  of  coal  a  year,  or  at  least 
15  per  cent  of  the  fuel.  A  covering  of  magnesia,  hair-felt,  or 
asbestos,  i  inch  thick,  would  in  five  years'  time  save  much  more 
than  its  cost,  which  is  25  cents  per  square  foot.  A  brick  arch 
over  the  boiler,  furnishing  a  conduit  for  flue- gases  on  their  way  to 
the  chimney  would  be  still  better.  Boilers  in  batteries,  if  placed 
back  to  back  instead  of  facing  one  .another,  would  also  reduce 
the  radiation  loss,  though  the  cost  of  stoking  would  be  increased. 

Flue-gases. — The  flue-gases  have  a  temperature  between  300° 
and  600°  F.,  being  higher  with  the  forced  draft  than  with  the  chim- 
ney draft.  In  order  to  give  place  for  the  air-supply  in  the  com- 
bustion of  the  coal,  these  flue- gases  are  removed  as  promptly  as 
possible.  Each  pound  of  coal  produces  gases  that  transport 
1500  or  more  B.T.U.  This  is  always  more  than  10  per  cent  of 
the  total  available  heat  in  the  fuel.  The  only  recoverable  por- 
tion of  it  is  that  which  the  economizer  may  utilize.  Smoke 
increases  the  heat  loss  by  that  which  is  latent  in  the  unconsumed 
combustible. 


p6  MANUAL  OF   MINING. 

The  Air-supply. — The  aggregate  of  the  above  sources  of  heat 
loss  is  at  least  15  per  cent  of  the  fuel  value,  and  is  more  frequently 
35  per  cent  in  the  ordinary  boilers. 

Economic  combustion  will  reduce  the  great  waste  of  heat  by 
providing  the  proper  supply  of  air,  the  perfect  burning  of  a  well- 
selected  fuel,  and  a  good  regulation  of  the  draft.  Conditions 
for  complete  combustion  require  that  the  particles  be  brought 
into  intimate  contact  with  the  requisite  amount  of  oxygen.  The 
amount  of  air-supply  being  the  chief  consideration,  each  pound 
of  coal  will  require  an  amount  of  air  at  least  equal  to  A  = 
i.52(C+3H—o.4O),  in  which  C,  H,  and  O  are  the  percentages 
of  carbon,  oxygen,  and  hydrogen. 

This  is  about  160  cu.  ft.  or  12  Ibs.  of  air.  More  than  this  is 
necessary  to  insure  intimate  contact  of  the  fuel  constituents  and 
the  air.  Hence  24  Ibs.  are  usually  given  for  each  pound  of  fuel 
with  the  chimney  draft  and  about  18  Ibs.  with  forced  draft.  Any 
excess  over  this  amount  chills  the  fire  and  will  only  carry  heat 
from  the  furnace  to  the  chimney.  With  less  air  the  carbon  will 
develop  only  4450  B.T.U.  per  pound,  instead  of  14,500  B.T.U., 
which  is  more  wasteful  than  smoke.  The  production  of  CO  in- 
stead of  CO2  results  from  a  deficiency  of  air;  the  fire  is  dull  and 
sluggish. 

An  occasional  analysis  of  the  flue-gases,  by  the  D'Orsat  ap- 
paratus, for  free  oxygen  and  CO,  will  reveal  the  working  of 
the  furnace.  The  former  indicates  air  excess,  the  latter  air 
deficiency. 

Calorific  Value  of  Fuels. — The  heating  power  of  a  fuel  depends 
upon  its  combustible  elements  and  the  amount  of  ash.  The  fuel 
should  be  bought  on  its  calorific  value  and  according  to  analysis. 
This  latter  may  be  the  ultimate  analysis,  giving  the  percentages 
of  carbon,  oxygen,  and  hydrogen  in  the  fuel,  whence  its  calorific 
value  is  P  =  i4,55oC+62,oooH— 54oo(O  +  N);  or  an  approximate 
analysis  giving  the  amount  of  volatile  matter  and  fixed  carbon  in 
the  fuel.  In  the  latter  case,  assume  M  to  be  the  percentage  of 

volatile  matter,  V,  compared  with  the  dry  combustible,  M  =  77-7^, 

~ 


POWER  GENERATION.  97 

and  A  a  coefficient  whose  values  are  given  below.     Whence  the 
calorific  power:  P^i^.jC+AM. 

A  =  252  for  M=   2  to  12;  4  =  169.2  for  M  =  30  to  35; 

A  =  216  for  If  =12  to  17;  4  =  144.0  for  Af  =  35  to  38; 

4  =  198  for  M  =17  to  24;  4  =  142.2  for  M  =  38  to  40; 

4  =  183  for  M  =  24  to  30;  4  =  136.8  for  M  =  40  to  50. 

The  heat  which  can  be  developed  from  a  pound  of  coal  varies 
between  11,000  B.T.U.  and  15,000  B.T.U.  As  one  pound  of 
water  requires  for  its  evaporation  and  conversion  to  steam  some- 
thing over  1 100  B.T.U.  it  follows  that  from  10  to  14  Ibs.  of  steam 
can  be  produced  for  each  pound  of  coal,  though  some  boilers 
realize  only  4  to  6  Ibs. 

Undoubtedly  the  semi-bituminous  coals  are  better  steam- 
producers  than  anthracite  coals,  the  amount  of  the  ash  being 
less.  But  it  is  usually  a  question  of  local  price  as  to  which  is 
preferable  in  a  given  case.  A  superior  coal  is  that  which  has 
about  85  per  cent  of  fixed  carbon  with  some  volatile  matter  to 
assist  in  the  firing.  A  steaming  coal  should  kindle  readily  and 
burn  steadily  without  clinkers. 

Slack  is  nearly  equal  to  coal  in  its  calorific  value,  but  has 
usually  too  much  refuse  to  be  acceptable,  except  where  the  trans- 
portation is  cheaper.  The  size  of  the  fuel  is  selected  according 
to  the  relative  grate  openings.  The  smaller  coals  and  slack  are 
cheaper  than  the  larger  size  of  the  same  coals,  but  the  amount  of 
ash  in  the  former  increases  the  labor  of  handling. 

The  Rate  of  Combustion. — The  amount  of  coal  that  can  be 
burned  depends  upon  the  intensity  of  the  draft  and  the  air- supply. 
The  thickness  of  the  bed  should  be  as  small  as  possible  without 
requiring  excessive  watchfulness.  A  5 -inch  bed  of  anthracite 
pea  coal  or  an  8-inch  bed  of  lump  coal  is  good  practice.  Only 
12  Ibs.  of  anthracite  or  15  Ibs.  of  bituminous  coal  can  be  burned 
per  hour  per  square  foot  of  grate  area  by  natural  draft  with  eco- 
nomical results.  The  results  from  the  boilers  depend  more  upon 
the  character  of  firing  than  on  any  other  single  element. 

The  Methods  of  Hand  Firing  usually  adopted  in  practice 
with  ordinary  grates  are  the  "spreading"  method,  in  which  a 


98  MANUAL  OF  MINING. 

thin  layer  of  coal  is  worked  over  the  entire  area  from  the  bridge 
towards  the  front;  the  "alternating"  method,  in  which  the  fresh 
coal  is  charged  alternately  on  either  side  of  the  fire-box;  and  the 
"coking''  system,  for  bituminous  coal  only,  that  charges  it  on  the 
dead-  plate  at  the  front  of  the  fire  and  pushes  it  back,  when  coked, 
into  the  grate  to  make  room  for  a  new  charge.  This  is  advan- 
tageous with  all  long-flame  gaseous  coals.  Two  stokers  and  two 
engineers  can  fire  35  tons  of  coal  per  week  under  the  boilers  during 
24  hours.  The  limit  for  one  man  firing  boiler  and  running  the 
engine  is  about  10  tons  of  coal  per  week  for  a  50-H.P.  engine. 

Stokers. — Mechanical  systems  of  feeding  hard  coal  are  employed 
to  advantage.  The  Steam  Users'  Association  of  Boston  reports 
that  in  all  plants  stokers  may  prevent  smoke,  save  coal,  and 
reduce  labor — as  much  as  30  per  cent  in  big  plants  employing 
also  coal-handling  machinery.  Mechanical  stoking  reduces  the 
smoke  nuisance  by  furnishing  a  continuous  charging.  Stokers, 
however,  cut  down  the  capacity  slightly,  though  they  respond  to 
a  sudden  demand  for  steam  as  well  as  hand  firing.  The  amount 
of  repairs  is  not  excessive. 

The  stokers  are  travelling  grates  or  moving  bars  fed  from  a 
hopper  continuously  at  the  front  and  carried  by  power  toward 
the  rear  at  a  slow  rate  such  as  to  insure  a  complete  burning  of  the 
fuel  before  it  reaches  the  rear  of  the  fire-grate  where  it  is  dumped. 
Fine  sizes  of  coal  can  be  utilized  and  little  power  is  required. 
Undoubtedly  they  afford  a  betterment  of  evaporative  results,  but 
it  is  a  question  whether  the  same  amount  of  improvements  can- 
not be  effected  in  other  directions. 

Fuel  Consumption. — The  consumption  of  fuel  per  hour 
averages  3!  Ibs.  per  horse-power  with  a  common  boiler  and 
ordinary  engine.  The  very  best  results  which  have  yet  been 
obtained  in  practice  require  1.5  Ibs.  of  good  coal  to  produce  a 
horse-power  hour. 

The  question  of  fuel  economy  is  an  exceedingly  important 
one,  whether  one  considers  the  production  of  power  at  the  coal- 
mine, where  the  fuel  is  cheaper,  or  at  the  metal-mine,  where  it  is 
very  dear.  Any  economy  effected  in  this  direction  represents  a 


POWER  GENERATION.  99 

corresponding  gain.  The  bituminous  collieries  of  the  United 
States  consume  between  1.5  per  cent  and  2  per  cent  of  their  total 
output  in  providing  power  for  pumping  and  water.  The  anthra- 
cite coal-mines  of  the  United  States  are  estimated  as  consuming 
from  8  per  cent  to  10  per  cent  of  their  total  production  for  the 
same  purpose. 

A  cord  of  well-dried  spruce  is  capable  of  driving  a  common 
slide-valve  hoister  about  320  hourly  horse-powers;  a  ton  of 
lignite,  470;  a  barrel  of  petroleum,  170;  a  ton  of  anthracite,  650. 

Weathering  Coals. — Fuels  should  be  properly  housed  from 
the  weather,  and  not  piled  out  of  doors,  to  prevent  absorption 
of  moisture,  the  evaporation  of  which  causes  deterioration. 
Freshly  mined  coal  absorbs  three  times  its  volume  of  gas  or  air, 
and  this  absorption  produces  a  slight  decomposition  that  reduces 
its  heating  power.  Moreover,  spontaneous  combustion  may 
develop  from  a  prolonged  oxidation,  which  can  be  prevented 
only  by  a  thorough  ventilation  of  the  coal  and  its  protection 
from  the  elements.  The  practice  of  sprinkling  the  coal  arrests 
the  action  temporarily,  but  deterioration  soon  begins,  which  would 
be  facilitated  by  the  presence  of  fine  coal.  It  is  a  good  practice  to 
have  a  storage  for  500  Ibs.  of  coal  for  each  indicated  horse- power 
of  engine  plant. 

Liquid  Fuels. — Petroleum  has  an  approximate  analysis  of 
carbon  85  per  cent,  hydrogen  13  per  cent,  and  oxygen  i  per  cent, 
and  a  calorific  value  of  about  20,000  heat-units  per  pound.  This 
fuel  can  be  injected  into  the  fire-box  from  tanks  by  a  nozzle 
with  the  aid  of  compressed  air.  This  blast  induces  a  current, 
atomizes  the  oil,  and  delivers  a  spray  with  an  excess  of  oxygen 
which  permits  the  production  of  an  intense  flame.  No  grate  is 
required.  One  pound  of  oil  equals  2.18  Ibs.  of  coal  in  calorific 
value.  Where  conditions  favor  it,  gas  from  coke-ovens  can  be 
employed  to  advantage  under  a  boiler. 

Draft. — A  draft  is  necessary  to  carry  off  the  products  of  com- 
bustion and  to  admit  a  fresh  supply  of  air  to  the  fire.  The  force 
of  the  draft  is  measured  by  a  water-gauge  which  varies  between 
r  inch  and  2  inches.  It  may  be  either  natural  or  forced. 


ioo  MANUAL  OF  MINING. 

Chimneys.  —  Natural  draft  is  procured  by  a  chimney  of  brick 
or  of  sheet-steel  tubing.  It  depends  upon  (he  relative  densities 
of  the  hot  gases  inside  the  chimney  and  the  external  air.  The 
difference  in  the  temperatures  and  the  height  of  the  stack  will 
determine  the  velocity  of  the  flow  and  the  pressure  of  the  draft. 
The  inside  area  determines  the  volume  of  air  which  can  be  car- 
ried. This  is  known  from  the  total  hourly  consumption  of  fuel, 
F,  which  is  based  on  5  Ibs.  per  horse-power.  The  desired  horse- 
power being  known,  the  effective  area,  E,  may  be  determined 
for  a  known  height  of  chimney,  H. 

0.06257? 

=~~ 


The  effective  area  E  js  less  than  the  actual  area  by  the  amount 
indicated,  E=A-  o.6\/Z. 

The  commercial  horse-power  of  chimney  is  ^.^EVH. 

The  first  cost  of  a  brick  chimney  is  greater  than  that  of  a 
sheet-steel  stack,  but  its  stability  is  greater  and  its  repairs  are  less. 
By  the  use  of  a  damper  regulated  by  hand  or  automatically  by 
steam,  any  lower  pressure  than  the  maximum  draft  for  which  it 
is  designed  may  be  maintained  in  the  latter  arrangement.  A 
small  diaphragm  raises  the  lever  and  closes  the  damper  when 
the  steam  pressure  in  the  boiler  exceeds  a  certain  desired  limit. 

Blowers.  —  Forced  draft  may  be  obtained  by  the  use  of  a 
blower  driven  by  the  engine  and  dynamo  or  belt.  By  it  air  is 
delivered  through  the  pipes  into  the  fireplace  at  a  pressure  of 
perhaps  10  Ibs.  above  the  atmosphere.  The  steam-jet  may  also 
be  employed,  acting  like  an  injector;  though  it  is  not  an  eco- 
nomical device,  it  softens  the  clinkers  and  aids  combustion. 
A  steam-blower  can  be  built  composed  of  a  horizontal  ring  with 
holes  punctured  in  it.  Steam  from  the  boiler  is  discharged 
through  the  openings  and  creates  a  current  that  will  increase 
the  draft  and  thereby  the  capacity  of  the  boiler  20  per  cent  to 
30  per  cent.  The  ring  is  made  of  i^-inch  pipe  12  inches  in  diam- 
eter punctured  with  eighteen  £-inch  holes  for  a  6o-horse-power 
boiler.  Such  a  blower  costs  90  cents  an  hour  for  operation. 


POWER  GENERATION.  IOI 

Exhaust- fans  may  be  so  placed  as  to  draw  the  gases  out  of 
the  furnace  and  deliver  them  into  short  stack  to  aid  draft. 

The  chimney  has  a  higher  first  cost  than  any  of  these  sys- 
tems which  require  short  stack  only,  but  is  cheaper  to  maintain. 
It  is,  however,  very  wasteful  of  heat,  since  the  flue-gases  must 
be  hot  in  order  to  produce  draft.  Nevertheless,  the  latter  adapts 
the  boiler  to  a  wide  range  of  power,  for  the  fire  can  be  made  more 
intense. 

Boiler  Installation. — The  boilers  should  be  on  a  level  below 
the  engine-room  and  set  as  high  up  as  is  consistent  with  the 
fireman's  duties.  The  coal  should  be  stored  above  them  and 
gravitation  availed  of  for  feeding  the  boilers  and  for  removing 
ashes.  This  would  be  an  ideal  condition  as  to  labor.  The 
cost  of  power  production  is  not  always  a  matter  only  of  fuel 
selection  and  economy  of  combustion,  for  the  cost  of  handling 
is  inseparably  connected  with  it.  Generally  speaking,  the  annual 
coal  consumption  equals  the  initial  cost  of  the  boiler  plant,  and 
economy  in  labor  may  easily  be  effected  equal  to  10  per  cent  of 
the  initial  cost. 

With  the  pipes  hung  to  drain  towards  the  boiler,  covered 
with  a  layer  of  asbestos,  magnesia,  or  felt  covering,  with  a  steam- 
trap  properly  placed,  a  supply  of  dry  steam  will  be  delivered 
to  the  engine.  Each  square  foot  of  exposed  surface  of  pipe 
radiates  1.8  B.T.U.  per  hour  for  each  degree  Fahrenheit  differ- 
ence in  temperature  between  the  steam  and  the  external  air. 
Thus  a  4- inch  pipe  carrying  steam  of  90  Ibs.  pressure  (320°  F.) 
40  feet  long  radiates  an  amount  of  heat  equivalent  to  17  Ibs. 
steam  from  i2o°F.  to  90  Ibs.  absolute  pressure.  Hence  a  con- 
siderable loss  of  power  ensues  if  pipes  be  uncovered  Every 
284  feet  of  4- inch  pipe  covered  with  i-inch  layer  saves  a  horse- 
power. 

As  the  allowable  velocity  of  steam  in  pipes  is  ordinarily  50  feet 
per  second,  it  is  necessary  to  have  a  line  of  pipes  ample  in  area  to 
furnish  a  copious  supply  of  steam  with  but  small  loss.  A  4-inch 
pipe  will  discharge  115  Ibs.  of  steam  per  minute,  the  initial  pres- 
sure being  100  Ibs.  gauge,  or  it  will  deliver  109  Ibs.  of  steam  at 


102  MANUAL  OF  MINING. 

80  Ibs.  gauge.  A  2-inch  pipe  will  deliver  30  Ibs.  and  24  Ibs. 
of  steam  respectively.  As  it  requires  about  25  Ibs.  of  steam 
to  supply  one  horse-power  per  hour,  a  2-inch  pipe  will  serve 
for  about  60  horse-powers. 

The  maximum  absolute  steam  pressures  for  successful  opera- 
tion are  ordinarily  as  follows:  For  simple  non-condensing 
engines,  60  Ibs.  gauge  pressure;  for  compound  non-condensing, 
loo  Ibs.  gauge;  for  simple  condensing,  90  Ibs.  gauge;  and  com- 
pound condensing,  130  Ibs.  gauge.  The  character  of  the  boiler 
will  largely  determine  the  maximum  pressure  available.  A 
multitubular  boiler  cannot  be  forced  beyond  100  Ibs.  pressure, 
while  the  water-tubular  boiler  may  be  carried  to  200  Ibs.  or  more. 

The  Steam-engines.  —  The  steam  will  be  conducted  to  the 
underground  pumps,  the  exhaust-fan  or  the  ventilating  steam- 
jet,  and  to  the  steam  end  of  all  compressors,  electric  generators, 
and  hoisters.  Its  behavior  in  the  cylinders  of  these  engines 
is  alike  in  all,  and  the  following  considerations  will  apply  to 
all. 

Classification  of  Engines.  —  The  engines  may  be  classified 
according  to  the  manner  in  which  they  utilize  the  elastic  force 
of  the  steam.  Direct-acting  pumps  and  other  engines  having  a 
constant  resistance  do  not  admit  of  a  change  of  pressure  within 
the  cylinder  of  the  engine;  but  where  the  resistance  varies,  or 
it  is  possible  to  provide  for  the  varying  resistance,  the  steam  is 
permitted  to  expand,  with  the  corresponding  reduction  of  pressure, 
with  marked  economy. 

The  Behavior  of  Steam. — Fig.  37  illustrates  the  behavior 
of  steam  inside  of  a  pump  in  which  the  length  of  the  diagram 
corresponds  to  the  length  of  the  cylinder,  and  the  height  of  the 
diagram  ihe  pressure  of  steam  at  various  points.  The  upper 
horizontal  line  represents  the  steam  pressure  admitted  to  the 
cylinder,  which,  remaining  constant,  will  be  straight.  The  atmos- 
pheric pressure  being  also  constant,  the  former  will  be  parallel 
to  it  and  at  a  distance  above  it  according  to  the  scale  on  which 
the  pressure  may  be  drawn.  The  pressure  of  the  atmosphere  being 
14.7  Ibs.  per  square  inch,  a  line  may  be  drawn  below  it  at  that 


POWER  GENERATION.  103 

scale  and  represents  the  line  of  absolute  zero  of  pressures.  OX 
is  that  line,  MN  the  atmosphere  line,  and  AB  the  line  of  con- 
stant steam  pressure.  When  the  piston  reaches  the  end  of  its 
stroke  the  exhaust-valve  is  opened  for  the  discharge  of  the  steam. 
The  pressure  falls  nearly  to  atmospheric  pressure.  On  the  return 
stroke  of  the  piston  the  steam  is  being  expelled  at  a  constant 
pressure,  represented  by  the  line  CD,  until  the  piston  is  at  the 
end  of  the  cylinder.  Steam  is  now  admitted.  The  pressure 
rises  and  the  line  DA  represents  such  increase  in  the  pressure. 

On  the  other  side  of  the  piston,  the  operations  ensue  in  the 
same  order,  the  line  CD,  representing  the  back  pressure,  on  one 
side  being  simultaneous  with  the  steam  pressure,  AB,  on  the  other 


FIG.  37. — Indicator  Card  from  a  Non-expansion  Steam-engine. 

side.  If  the  steam  be  exhausted  into  the  condenser,  the  line  CD, 
representing  the  back  pressure  upon  the  pump,  will  be  below  the 
atmospheric  line  MAT".  Its  length  above  the  zero  line,  OX,  is  an 
amount  representing  the  vacuum  pressure  in  the  condenser. 
This  line  would  then  be  at  about  5  Ibs.  as  a  minimum. 

If  the  steam  after  being  admitted  at  constant  pressure  for  a 
portion  of  the  stroke  is  cut  off,  the  piston  will  continue  to  move  in 
its  stroke,  but  with  a  continual  reducing  pressure  to  the  end.  Fig. 
38  will  represent  in  the  two  lines  AB  and  BI  the  behavior  of 
the  steam  during  admission  and  expansion.  The  back-pressure 
line,  CD,  will  remain  the  same  as  in  the  pump.  The  discharge 
of  the  steam  may  be  cut  off  at  some  point,  D,  on  the  return  stroke; 
the  steam  still  remaining  in  the  cylinder  will  be  compressed  and 
the  line  DE  will  indicate  the  increase  in  pressure  for  the  balance 
of  the  stroke.  When  admission  occurs  the  line  EA  will  rise 
more  or  less  promptly,  according  to  the  freedom  of  entry  of  steam. 
The  back-pressure  line  will  be  below  or  above  MN  according 


104  MANUAL  OF  MINING. 

to  whether  a  condenser  is  attached  or  not.  The  diagrams  repre- 
sent, therefore,  the  behavior  of  the  steam  at  all  periods  of  the 
piston  movements  in  the  cylinder,  and  the  area  enclosed  be- 
tween ABIC  DBA  measures  the  work  done  during  one  stroke. 
The  vertical  lines  represent  the  pressures  to  scale,  in  pounds  per 
square  inch  of  area  piston;  the  horizontal  line  represents  the 
length  of  stroke,  s.  The  area  divided  by  the  length  of  card  and 
multiplied  by  the  scale  of  vertical  pressures  equals  the  mean 
effective  pressure,  M.E.P.,  in  the  piston  during  the  entire  stroke. 
Diagrams  such  as  these  are  called  indicator  cards  and  measure 
the  indicated  horse-power  of  the  engine.  The  device  known 


102.fi* 


FlG.  38. — Indicator  Diagram  from  a  Non-condensing  Engine  or  the  High-pressure 
Cylinder  of  a  Compound  Engine. 

as  the  indicator  (Fig.  40)  can  be  procured,  by  which  such  cards 
of  steam  behavior  can  be  automatically  registered  by  the  engine 
while  at  work. 

In  order  to  obtain  the  most  economical  consumption  of 
steam,  the  lowest  point,  /,  of  the  expansion  line,  BI,  should  be 
as  near  to  the  back-pressure  line  as  possible.  The  limit,  however, 
is  3  Ibs.  gauge  in  non-condensing  engines  and  about  5  Ibs.  abso- 
lute when  condensers  receive  the  exhaust. 

The  expansion  line,  BI,  follows  practically  the  equation  of 
the  equilateral  hyperbola,  PV  =  constant;  if  the  cylinder  walls  be 
perfect  non-conductors,  the  expansion  line  would  be  represented 
by  the  adiabatic  equation  py1-135. 

The  most  desirable  point  of  cut-off  then  is  fixed,  the  conditions 
being  given.  Thus,  if  the  pressure  at  /be  18  Ibs.  absolute  and 
the  admission  pressure  be  90  Ibs.  absolute,  the  cut-off  will  be  at 
one-fifth  of  the  stroke.  The  ratio  of  expansion  is  then  5;  that 


POWER  GENERATION. 


is,  the  volume  of  steam  delivered  to  the  cylinder  up  to  point  of  cut-off 
is  expanded  to  five  times  this  volume  at  the  end  of  the  cylinder. 

The  Mean  Effective  Pressure. — By  the  following  equation 
M.E.P.  may  be  ascertained  for  a  given  admission  gauge  pres- 
sure, P,  and  a  back  pressure,  B,  in  pounds  absolute  per  square 
inch.  By  gauge  pressure  is  understood  the  pressure  above 
the  atmosphere  as  recorded  on  the  steam-gauge.  Absolute  pres- 
sure is  the  pressure  above  absolute  vacuum.  Atmospheric  pressure 
is  14.7  Ibs.  absolute.  />=C(P+  14.7)-^. 

The  value  for  C  is  obtained  from  the  following  table  for  given 
ratios  of  cut-off. 

VALUE  FOR  THE  CONSTANT,  WITH  VARIOUS  CUT-OFFS,  FOR  DETERMINING 
THE  M.E.P. 


. 

Ratio  of 

Ratio  of 

Percentage 
Cut-off 
Apparent. 

Expansion, 
No  Clear- 
ance. 

C. 

Expansion 
with  3 
Per  Cent 

C. 

Expansion 
with  7 
Per  Cent 

C. 

Clearance. 

Clearance. 

O.I 

10 

•33°2 

7.69 

•  4113 

5.88 

'47*5 

f 

8 

•385° 

6-45 

•4435 

5.128 

•5i37 

i 

6 

•4653 

5.102 

.5155 

4-23 

•5774 

i 

5 
4 
3 

•5219 
.5966 

•6995 

4.3478 
3-57J4 
2-754 

.5682 
.6368 
.7294 

3-7037 
3-125 
.481 

.6225 
.6861 
.7708 

0.4 

2-5 

.7664 

2-3255 

•7931 

.1276 

.8242 

| 

2 

.8465 

1.8867 

.8658 

•754 

•  8925 

| 

1.6 

.9180 

1-527 

.9310 

•439 

.9481 

I 

1-333 
1.25 

.9656 
•9785 

1.282 
1.2875 

•9743 
.9807 

•2195 
.149 

.9842 
.9914 

I 

1.  143 

.9918 

1.104 

.9941 

.058 

.9961 

For  all  practical  purposes  the  ratio  of  expansion  and  the 
point  of  cut-off  are  reciprocals  of  one  another.  The  clearance 
space  is  left  at  the  end  of  the  cylinder,  which  the  piston  never  fills 
and  in  which  steam  is  allowed  to  accumulate  and  act  as  a  cushion 
to  prevent  the  piston  from  striking  the  end  of  the  cylinder.  This 
clearance  space  contains  steam  which  is  ineffective  and  changes 
somewhat  the  ratio  of  expansion  for  a  given  cut-off.  The  clear- 
ance is  from  3  per  cent  of  the  stroke  in  the  better  classes  of 
engines  to  7  per  cent  in  the  simple  slide-valve  engines.  High- 
speed engines  may  have  a  still  larger  clearance. 

The  shorter  the  point  of   cut-off  the  greater  the  efficiency  of 


106  MANUAL  OF  MINING. 

the  engine.  The  limit,  however,  depends  upon  the  character  of 
the  admission-valve.  With  a  simple  slide-valve  engine  the  cut-off 
cannot  be  less  than  0.4  of  the  stroke.  With  the  Corliss  valves, 
or  some  forms  of  rider  cut-off  valves,  the  cut-off  may  be  very  much 
shorter,  though  it  rarely  is  less  than  one-fifth  in  a  single  cylinder. 

The  M.E.P.  and  the  area  of  the  card  in  Fig.  37  may  be  greater 
than  in  Fig.  38,  but  it  must  be  remembered  that  the  volume  of 
steam  used  in  the  former  case  is  equal  to  the  entire  volume  of 
the  cylinder,  and  in  the  latter  case  to  the  volume  up  to  the  point 
of  cut-off.  Hence  the  work  performed  per  pound  of  steam  in 
the  latter  case  will  be  greater  than  in  the  former. 

The  Use  of  the  Indicators. — An  indicator,  which  is  essentially 
a  stethoscope  to  the  engine,  reveals  the  character  of  the  steam 
expansion  compared  with  the  theoretical;  the  amount  of  water 
existing  in  the  cylinder  at  any  moment;  the  point  of  cut-off;  the 
setting  of  the  valves;  and  the  amount  of  resistance  to  exhaust. 
None  of  these  can  be  detected  from  the  behavior  of  the  engine. 
Leaks,  tight  stuffing-boxes,  loose  piston  packing,  or  a  priming 
of  the  admission  steam  are  likewise  revealed  by  the  indicator. 
The  machine  is  simple,  easily  attached,  and  requires  no  elaborate 
calculation,  and  by  the  aid  of  the  planimeter  (Fig.  39)  it  is 


FIG.  39. — Measuring  the  M.E.P.  of  an  Indicator  Card  by  Planimeter. 

possible  to  obtain  a  card  by  which  to  determine  the  indicated 
horse-power.     Fig.  40  illustrates  one  attached  to  the  engine. 

An  indicator  is  a  device  for  registering  the  behavior  of  the 
steam  within  the  engine  cylinder.     It  consists  of  a  drum  and  a 


POWER  GENERATION. 


107 


cylinder;  the  latter  has  a  plunger  whose  rod  carries  a  pencil  above 
it  and  is  connected  by  a  pipe  to  the  engine  cylinder.  When 
desired,  the  steam  may  be  admitted  from  the  engine  under  the 
plunger,  which  is  then  lifted  an  amount  depending  upon  the  pres- 
sure existing  in  the  engine  at  that  instant.  With  it  the  pencil 
rises,  and  each  change  of  steam  pressure  will  be  indicated  by  a 
corresponding  rise  or  fall  of  the  pencil.  To  control  the  amount 
of  lift  of  the  piston  and  its  pencil  a  spring  of  known  strength  is 
inserted  above  the  plunger,  and  thus  not  only  each  change  of  pres- 
sure but  also  the  amount  of  such  pressure  may  be  recorded. 
Each  rise  of  pressure  lifts  the  plunger  a  definite  amount  by  com- 
pressing the  spring  correspondingly.  A  drop  in  the  pressure  is 
indicated  by  an  extension  of  the  spring.  The  pencil  then  rises 
a  certain  amount  when  the  steam  pressure  in  the  engine  and  on 
the  plunger  is  100  Ibs.  per  square  inch.  It  will  rise  half  that  for 
a  steam  pressure  of  50  Ibs.,  etc. 


FIG.  40. — Attaching  an  Indicator. 

The  pencil  bears  against  and  records  upon  a  sheet  of  paper, 
about  6  inches  long,  which  is  slipped  on  a  drum  of  about  2  inches 
in  diameter.  A  large  wheel  attached  to  the  drum,  turned  by 
a  string  fastened  to  the  cross-head  of  the  engine,  pulls  the  string, 
turns  the  drum,  and  the  paper  moves  under  the  pencil.  As  the 
piston  travels  in  its  cylinder  the  pencil  rises  and  falls  simultane- 
ously with  it,  thus  furnishing  a  diagram  called  an  indicator  card. 
Various  devices  are  employed  to  reduce  the  long  piston  stroke 
to  a  card  length  of  about  4  inches. 


joS 


MANUAL  OF  MINING. 


The  plunger  springs  are  selected  according  to  the  boiler  pres- 
sure, so  that  the  height  of  lift  of  pencil  shall  not  exceed  2  inches^ 
thus  for  100  Ibs.  admission  pressure  a  number  50  spring  is  taken, 
which  means  that  steam  of  50  Ibs.  pressure  will  compress  that 
spring  an  amount  that  will  lift  the  pencil  i  inch;  each  change  of 
i  Ib.  steam  pressure  will  move, the  pencil  -^  of  an  inch.  The 
scale  of  the  spring  is  50.  On  the  diagram  drawn  by  such  a 
spring  the  vertical  lines  represent  the  pressure  at  the  rate  of  50 
Ibs.  per  inch  of  height,  and  the  horizontal  line  the  stroke  of 
the  piston  on  a  scale  equal  to  the  ratio  of  the  piston  stroke  divided 
by  the  card  length. 

Steam  Condensation  in  the  Cylinders — Remedies. — The  objec- 
tion to  a  short  cut-off  and  a  high  rate  of  expansion  lies  in  the  fact 
that  the  walls  of  the  cylinder  are  not  non-conductors,  which,  being 
comparatively  cool  at  the  end  of  the  stroke,  condense  some  of 
the  entering  steam  whose  energy  is  thus  lost.  The  greater  the 
ratio  of  expansion  the  larger  is  the  percentage  of  the  steam, 
condensed.  It  may  reach  as  much  as  50  per  cent  during  some 
portion  of  the  stroke  in  some  engines,  and  is  always  25  per  cent 
with  unprotected  cylinders.  This  is  reduced  by  wooden  lagging 
around  the  cylinders  or  steam  circulating  in  an  annular  space 
around  the  cylinder.  Another  remedy  consists  in  superheating  the 
steam  after  leaving  the  boiler  by  passing  it  through  the  pipes 
heated  by  flue-gases  before  delivering  to  the  engine. 

A  better  remedy  is  to  divide  the  expansion  into  several  stages 
by  compounding   the    cylinders. 
In  this  the   steam,  after  partial 
expansion  in  one  cylinder,  is  de- 
15   livered   to  a  second,   and  there 
j    expanded  to  the  tower  limit  de- 
termined by  the  pressure  in  the 
condenser  or  the  atmosphere  (Fig. 
FIG.  41.— Combined  Indicator  Cards.   4i)-    The  action  in  the  combined 

cylinders   is   the   same   as   in    a 
single  cylinder  with  the  same  ratio  of  expansion. 

The  Compound  Engine. — These  cylinders  may  be  connected 


POWER  GENERATION. 


109 


on  the  same  piston-rod  and  have  a  common  stroke  (tandem),  or 
are  side  by  side  on  different  rods  (cross- compound)  and  move 
in  opposite  directions.  If,  however,  the  motion  of  the  pistons 
is  to  be  at  some  other  phase,  a  receiver  is  inserted  between  the 
two  cylinders,  the  exhaust  from  the  high- pressure  accumulating  in 
it  before  being  discharged  into  the  low-pressure  cylinder. 


Steam  Exhausting 
into.L.P.  Steam  Chest     . 

{Effective  Pressure)    ti. 
£pace  A.  A.  A.  A.  is  a  Balanced  Pressure, 
between  the  two  Piston  Heads, 
tTandem  CompoitntLt'ofltaa 

FIG.  42. — The  Section  of  a  Compound  Cylinder. 

The  compound  engine  economizes  steam  by  permitting  a 
larger  ratio  of  expansion,  having  smaller  clearance  volumes  and 
far  less  condensation  in  the  cylinders.  The  cost  of  the  com- 
pound is,  however,  higher  than  the  simple,  its  bulk  and  weight 
are  greater,  but  its  steam  consumption  is  very  much  less. 

Governors. — To  control  the  engine  and  to  economize  steam 
by  supplying  to  the  cylinder  only  the  amount  that  is  necessary 
to  do  the  work,  two  forms  of  gov- 
ernors are  employed.  The  steam 
pressure  may  be  reduced  before  ad- 
mission to  the  cylinder  by  closing 
the  valve  in  the  steam-pipe;  or  an 
equally  automatic  regulation  may 

be   had  by   reducing   the   volume  of  FIG.  43.— Indicator  Diagrams  from  a 

the  steam  delivered  to  the  cylinder          Throttling  steam-engine. 
by  hastening  the  point  of  cut-off. 

In  Fig.  43  are  the  diagrams  obtained  successively  in  the 
throttling-engine  as  the  steam-supply  is  reduced.  Fig.  44 


HO  MANUAL  OF  MINING. 

shows  the  diagrams  from  a  cut-off  engine.  The  throttling  is 
obtained  by  a  pendulum  governor  connected  with  the  driving- 
shaft  of  the  engine,  which  closes  the  valve  when  the  rate  of  revo- 
lution exceeds  a  certain  limit.  The  cut-off  is  operated  by  some 
form  of  gearing  connected  with  the  shaft,  and  when  the  rate  is 
too  high  reduces  the  travel  of  the  valve.  The  throttling-engine 
is  cheaper  and  suffers  less  from  cylinder  condensation  than  a 
cut-off  engine,  but  does  not  receive  so  sensitive  or  so  uniform  a 
regulation.  All  engines  have  an  independent  throttling  control 
for  the  engineer. 

Automatic  Cut-off. — The  mechanism  'for  shutting  off  the 
steam-supply  before  the  piston  reaches  the  end  of  its  stroke  may 
be  fixed  in  position  or  variable,  automatic,  or  adjustable.  The 
fixed  cut-off  is  intended  for  constant  duty,  the  governor  being 
set  to  shut  off  the  steam-supply  when  the  desired  point  is  reached. 
The  variable  cut-off  is  needed  for  a  varying  load  during  suc- 
cessive strokes.  In  the  high-speed 
engine  the  shaft-governor  varies  the 
length  of  travel  of  the  slide-valve, 
and  thereby  hastens  cut-off  when 
the  load  is  light.  In  the  Corliss 
^*-i-  •  engine  the  governor  operates  di- 

FIG.  44  .—Indicator  Diagrams  from  rectly  upon    an  arm  which  varies 

an  Automatic  Cutoff  Engine.        ^  ^  Qf  ^  Qff  foy  changjng  ^ 

tripping  position.  The  latter  two  types  are  automatic  in  their 
action,  the  Meyer  being  adjustable.  Its  rider-valves  (Fig.  44) 
can  be  set  at  each  end  at  any  time  to  the  desired  cut-off  without 
affecting  the  movement  of  the  lower  valve  which  determines  the 
duration  of  the  exhaust.  The  Corliss  valves  are  also  adjustable 
by  changing  the  length  of  the  rods  from  the  wrist- plate  to  valve- 
spindles. 

Throttling  governors  are  usually  employed  on  all  low-speed 
engines  and  shaft-governors  on  high-speed  engines.  Engines 
arranged  to  run  "over"  are  smoother  than  those  turning  "under." 
Facing  the  engine  broadside  with  cylinder  on  the  left  and  crank- 
shaft at  the  right,  right-handed  revolution  is  called  "over." 


POWER  GENERATION.  in 

The  Sliding  Steam-valves. — The  valves  for  regulating  the 
admission  and  exhaust  of  the  steam  are  of  different  types.  Sim- 
ple engines  have  one  slide-valve  to  perform  all  the  functions  for 
both  ends.  Such  a  valve  admits  of  a  cut-off  not  shorter  than 
0.4,  and  wastes  considerable  power  in  supplying  the  steam.  The 
sliding  piston-valve  (Fig.  144)  is  preferable,  because  it  is 
balanced  and  requires  very  little  power  to  move  it,  but  it  does 
not  give  a  sharp  cut-off.  The  inability  of  either  of  these  to  alter 
the  cut-off  (B,  Fig.  38)  without  also  affecting  the  degree  of 
compression,  DE,  makes  the  Meyer  cut-off  valve  (Fig.  124)  a 
better  controller.  Separate  eccentrics  drive  the  main  slide- 
valve  and  the  rider-valve,  the  latter  to  control  the  point  of  cut- 
off only,  while  the  main  valve  regulates  the  exhaust  and  admits 
steam  when  the  rider  uncovers  its  port. 

The  Corliss  Valve  consists  of  four  rotary  valves,  each  with  a 
single  function  to  perform.  One  admission  rotary  valve  and 
one  exhaust  rotary  valve  serve  for  each  end  of  the  cylinder.  They 
are  all  moved  by  rods  from  a  wrist-plate,  the  admission-valve 
being  provided  with  a  releasing  trip-gear  to  make  a  sharp  cut- 
off (Fig.  45). 


FlG.  45. — Section  of  Corliss  Engine,  showing  valves  in  position  for  admitting 
steam  at  the  left. 


Fig.  126  illustrates  the  Corliss  engine  and  valve  in  section 
on  the  left-hand  air-cylinder. 

The  Reversing-link. — The  reversal  of  the  direction  of  motion 
<jf  the  engine  is  accomplished  by  one  of  two  methods:  engines 
supplied  with  shaft-governors  may  be  reversed  in  direction  by 
changing  the  position  of  an  eccentric  on  the  shaft  and  thus  mov- 
ing the  valve  to  correspond;  and  slow-speed  engines  by  the 


*I2  MANUAL  OF  MINING. 

Stevenson  link  motion  with  two  eccentrics  (Fig.  46).  Motion  is 
imparted  to  the  steam- valve  by  a  rod  connected  to  an  eccentric, 
on  the  crank-shaft.  The  eccentric  is  placed  at  an  angle  of  some- 
thing more  than  90°  in  advance  of  the  crank-pin  position.  When 
a  reversal  of  motion  is  desired  the  single  eccentric  is  turned  to 
an  angle  of  more  than  90°  in  advance  on  the  other  side  of  the 
crank-pin.  The  Stevenson  link  arc  has  each  end  connected 
with  an  eccentric  When  the  arc  is  raised  or  lowered  till  the 
block  to  which  the  valve  is  attached  is  at  the  extreme  end  of 
the  arc,  the  valve. will  have  a  maximum  travel.  In  position  a 


FIG.  46. — The  Reversing  Link. 

(Fig.  46)  the  block  gives  a  forward  motion,  and  when  raised  to 
the  position  b  a  backward  motion,  to  the  valve.  In  its  central 
position  the  block  is  acted  upon  by  both  eccentrics  and  only  a 
slight  motion  is  imparted  to  the  valve. 

The  Condenser  is  a  receiver  fitted  with  tubes  of  copper 
through  which  the  steam  may  flow  from  engine  to  a  tank  or  a 
pump  before  delivery  into  the  boiler.  Around  the  tubes  large 
volumes  of  cold  water  circulate,  condensing  the  steam  and 
creating  in  the  tubes  a  vacuum,  which  is  further  maintained  by 
the  air-pump.  The  vacuum  obtained  in  practice  is  about  25 
inches  of  mercury  column  below  the  atmosphere,  or  an  absolute 
pressure  of  2.5  Ibs.  per  square  inch.  This  addition  to  a  plant 
economizes  power  by  reducing  the  back  pressure,  increasing 
correspondingly  the  net  effective  pressure,  as  may  be  seen  by 
comparing  the  indicator  cards  (Figs.  38  and  47),  by  nearly  15  Ibs. 
per  square  inch.  This  is  the  quickest  and  simplest  means  of 


POWER  GENERATION.  113 

increasing  the  power  of  a  given  engine  and  boiler  plant.  The 
saving  in  steam  consumption  is  indicated  for  Corliss  engines 
in  a  later  table. 


FlG.  47. — Indicator  Diagram    from    a  Condensing  Engine  or  the  Low-pressure 
Cylinder  of  a  Compound  Engine. 

For  every  100  feet  of  piston  speed  the  condenser  will  save 
1.03  H.P.  in  an  engine  whose  cylinder  is  6  inches  in  diameter; 
with  a  lo-inch  cylinder,  2.86  H.P.  is  saved  per  100  feet  of  veocity; 
with  a  i6-inch  cylinder,  7.31  H.P.;  and  for  20  inches  diameter, 
11.42  H.P.  Thus  an  engine  with  io"Xi6"  cylinder  making 
80  r.p.m.  (213.33  feet  piston  speed)  saves  6.1  H.P. 

Influence  of  Clearance. — The  clearance  space  at  each  end  of 
a  cylinder  prevents  the  piston  striking  the  heads  and  provides  a 
space  in  which  steam  accumulates  to  form  a  cushion  and  assist 
in  the  reversal  of  piston  movement.  A  large  clearance  is  not 
economical.  Hence  high-speed  and  slide-valve  engines  are  less 
economical  than  the  Corliss.  A  large  clearance  implies  a  small 
degree  of  compression  in  the  cushion  and  less  back  pressure. 

The  most  economical  mean  effective  pressure  is  that  corre- 
sponding to  a  minimum  steam  consumption.  When  the  steam 
supply  is  cut  off  at  a  point  of  stroke  that  will  permit  it  to  expand 
to  the  pressure  of  the  condenser,  a  complete  expansion  is  obtained. 
The  more  complete  the  expansion  the  less  will  be  the  mean  effec- 
tive pressure.  With  a  low  M.E.P.  a  higher  economy  is  attained 
than  from  a  similar  engine  with  a  high  M.E.P.  A  low  pressure 
means  large  cylinders  and  excessive  condensation.  The  degree 
of  expansion  and  the  number  of  cylinders  are  fixed  by  the  ratio 
of  the  allowable  boiler  pressure  to  that  of  the  exhaust  into  con- 
denser or  atmosphere. 

An  engine  with  a  fixed  load  is  operating  at  its  maximum  effi- 
ciency when  driving  its  calculated  load.  If  the  load  varies,  the 


JI4  MANUAL  OF  MINING. 

-engine  must  have  a  capacity  for  the  maximum  demand.  If  this 
load  is  casual  and  seldom  demanded,  the  engine  may  be  then 
forced.  Its  size  will  be  determined  by  the  demand  of  the  average 
load,  in  supplying  which  it  may  attain  its  highest  efficiency. 

With  a  single- valve  simple  engine  the  most  economical  M.E.P. 
is  about  one  half  of  the  boiler  pressure,  which  would  represent  a 
steam  consumption  of  about  28  Ibs.  and  a  M.E.P.  of  50  Ibs.  per 
square  inch.  A  simple  Corliss  has  its  maximum  economy  at  a 
M.E.P.  of  35  Ibs.  pressure  and  a  Corliss  compound  at  26  Ibs. 
The  steam  consumption  would  be  25.3  and  19  Ibs.  respectively. 

Selection  of  Engine. — Generally  speaking,  any  expenditure  for 
a  plant  is  warranted  which  will  show  a  saving  more  than  equal  to 
the  interest  on  the  expenditure,  the  depreciation  and  repairs  of 
the  plant.  But  this  balance  is  not  easily  obtained.  An  economic 
engine  is  one  with  a  low  M.E.P.,  large  cylinders,  and  costly  at- 
tachments. But  its  initial  cost  may  be  prohibitive  at  the  time  of 
Its  installation.  If  the  fuel  is  cheap  and  the  amount  of  repairs 
not  known,  a  high-pressure  plant  may  not  be  desirable,  particu- 
larly as  its  piping  system  usually  gives  trouble  and  the  wear  and 
leakages  are  large.  Nevertheless  the  saving  resulting  from  its  use 
generally  more  than  offsets  the  increased  cost.  The  smaller 
boiler  capacity  and  the  diminishment  of  its  cost  should  be  con- 
sidered with  the  other  items  of  saving. 

The  Speed  of  Engines. — Engines  are  classified  according  to 
their  piston  speed,  as  slow-speed,  averaging  100  feet  per  minute, 
and  high-speed,  making  nearly  900  feet  of  travel  per  minute. 
This  value  is  the  product  of  the  number  of  revolutions  and  the 
piston  stroke — two  independent  variables. 

The  Corliss  engine  is  limited  to  about  60  r.p.m.,  and  electric 
engines  for  lighting,  etc.,  make  about  400  revolutions.  For 
mining  purposes  the  rate  rarely  exceeds  100  r.p.m.  The  high 
rotary  speed  is  not  economical,  because  of  the  large  clearance 
at  each  end  of  the  cylinder.  The  excellent  workmanship  and 
the  great  care  which  the  high-speed  engines  require  make  their 
first  cost  and  subsequent  maintenance  large.  They  have  less 
steam  condensation  than  slow  engines. 


POWER   GENERATION.  115 

The  work  of  the  piston  is  equal  to  the  continued  product  of 
the  mean  "effective  pressure  upon  It,  the  stroke  in  .feet,  its  area  in 
square  inches,  and  the  number  of  strokes  If  the  speed  is  fixed 
for  a  given  engine,  the  area  may  be  designed  proportionately  to 
s,  the  mean  effective  pressure  being  altered  at  will  (by  changing 
the  initial  boiler  pressure  or  the  cut-off). 

The  Horse-power  of  an  Engine. — The  theoretical  horse-power 
is  the  product  of  the  work  of  the  piston  and  the  number  of  strokes 
divided  by  33,000.  The  area  of  the  -piston-rod  on  the  crank  side 
is  neglected,  as  also  the  losses  from  friction  of  the  engine  parts, 
steam  condensation  in  the  cylinder,  clearance,  and  compression. 
The  value  for  the  effective  pressure,  p,  is  obtained  from  the 
formula  on  a  previous  page. 

The  indicated  horse-power  (I.H.P.)  is  the  actual  work  of 
the  steam  as  revealed  by  the  indicator  card;  the  mean  effective 
pressure  is  obtained  (Fig.  39)  by  the  card.  It  is  less  than 
p,  obtained  above  by  a  coefficient,  E,  whose  value  is  less  than 
unity. 

The  "brake"  horse-power  (B.H.P.)  is  a  certain  fractional 
part,  m,  of  the  indicated,  which  allows  for  the  mechanical  losses 
and  for  intermittence  in  running.  The  value  for  the  modulus 
varies  from  0.60  in  a  very  poor  engine  to  0.90  in  one  in  good 
working  order.  A  fair  average  for  the  mechanical  efficiency, 
m,  of  the  engine  ratio  between  the  brake  horse-power  and  the 
indicator  horse-power  is  0.8. 

Let  s  =the  length  of  the  piston  stroke  in  inches; 
&=the  diameter  of  cylinders  in  inches; 
N  =the  number  of  revolutions  per  minute; 
o.i66./Vs=the  piston  speed,  feet  per  minute; 
c=the  number  of  cylinders; 
/=the  coefficient  of  friction; 

^=mean  effective  steam  pressure,  pounds  per  square  inch; 
P=  absolute  steam  pressure  entering  the  cylinders; 
r  =  ratio  of  expansion  or  per  cent  of  cut-off; 
B  =  absolute  back  pressure  of  the  steam  during  exhaust; 
I?=horse-power  developed  by  the  engine. 


MANUAL  OF  MINING. 


Then 

H  =  (o.ooooo3g66csk2N)pm.  » 

In  the  following  table  are  calculated,  for  certain  sizes  of 
cylinders  and  given  piston  speeds,  the  values  for  the  parenthetical 
expression. 

THE  HORSE-POWER  FOR  EACH  POUND  OF   MEAN   EFFECTIVE  PRESSURE  PER 
SQUARE  INCH  OF  PISTON  AREA. 
The  value  for  o.ooooo3966s£2./V. 


Diam- 

Average Speed  of  Piston  in  Feet  per  Minute. 

eter  of 

Cylin- 

der. 

240 

300 

350 

400 

450 

Soo 

550 

600 

650 

750 

4 

.091 

.114 

•133 

.162 

.171 

5 

.144 

.18 

.21 

.24 

•27 

6 

.205 

.256 

.299 

•342 

.385 

.428 

.471 

7 

.279 

•  348 

.408 

.466 

•  524 

.583 

.641 

8 

.365 

.456 

•532 

.608 

.685 

.761 

.837 

.912 

9 

.462 

•577 

.674 

.770 

.866 

.963 

1.059 

-154 

10 

•571 

.714 

.833 

•952 

1.071 

1.190 

1.309 

.428 

i-547 

I-785 

it 

.691 

.864 

I.OOS 

I.I52 

1.296 

1.44 

1.584 

.728 

1.872 

2.  1  60 

12 

.820 

1.025 

I  •  195 

1.366 

1.540 

1.708 

1.  880 

.050 

2.222 

2.564 

14 

1.119 

1.398 

1.631 

1.864 

2.097 

2-331 

2.564 

•797 

3.029 

3-495 

15 

1.285 

i.  606 

1.873 

2.I3I 

2.409 

2.677 

2-945 

3.212 

3-479 

4.004 

16 

1.461 

1.827 

2.I3I 

2.436 

2.741 

3-°45 

3-349 

3-654 

3.958 

4-567 

18 

1.849 

2.312 

2.697 

3-083 

3.468 

3.854 

4.239 

4.624 

5.009 

5-/8 

20 

2.292 

2-855 

3-331 

3-807 

4.285 

4-759 

5-234 

5-73i 

6.186 

7-I38 

23 

3.021 

3-776 

4.104 

5-°35 

5.664 

6.294 

6.923 

7-552 

8.181 

9-44 

EXAMPLES. — i.  Required  the  horse-power  of  an  engine  6"X  10"  with  a  pis- 
ton spaed  of  300  feet  per  minute  (183  revolutions)  with  M.E.P.  of  40  Ibs. 
According  to  the  table,  for  k  =  6  inches  and  the  piston  speed  of  300  feet,  a, 
constant  of  0.256  is  found.  For  40  Ibs.  pressure  and  w=o.8o  the  horse- 
power, #"=0.256X40X0.80=8.19. 

2.  Required  the  horse-power  of  a  duplex  engine  io"Xi4.4"  making  100 
r.p.m.  and  having  a  cut-off  of  \  with  7  per  cent  clearance  and  initial  admis- 
sion pressure  of  100  Ibs.  absolute. 

Let  m=o.8o  and  back  pressure  18  Ibs.  according  to  table  on  page  105. 
Then  />= 0.6861(100) -18=50,64.  C=o.686i. 

According  to  the  above  table  the  constant  for  ic-inch  cylinder  and  240 
feet  piston  speed  is  0.5 71.  Whence  #"=0.571X50.64X0.80=23.13  for  each 
cylinder  and  46.16  for  the  duplex  engine. 

Usually  there  is  an  approximate  ratio  between  the  diameter,  k, 
and  the  length,  s,  of  a  cylinder.  .In  a  long- stroke  engine  s  =  2k  to 
2.56.  A  short-stroke  engine  has  a  ratio  s=k  to  s  =  i.2k. 


POWER  GENERATION.  117 

The  Steam  Consumption  of  an  Engine. — A  one-horse-power 
engine  exercised  through  one  hour,  equals  1,980,000  ft.-lbs.;  one 
B.T.U.  =778  ft.  Ibs.;  hence  one  horse-power  hour  corresponds 
to  2545  B.T.U.  A  pound  of  steam  represents  about  noo  B.T.U. 
above  feed-water  temperature.  Hence  one  horse-power  should 
be  obtained  during  one  hour  from  2.3  Ibs.  of  steam.  The  very 
best  on  record  is  12.2  Ibs.  per  I.H.P.,  and  for  a  stationary  engine 
of  the  Corliss  type,  compound,  about  18  Ibs.  Hence  all  steam  con- 
sumption above  12.2  Ibs.  may  be  regarded  as  preventable  waste. 

The  following  table  shows  the  consumption  of  steam  per 
hourly  horse-power  with  non-condensing  and  condensing  engines 
at  various  rates  of  cut-off  and  different  boiler  pressures.  An 
allowance  is  here  made  for  ordinary  cylinder  losses.  The  slide- 
valve  engine  would  show  a  steam  consumption  under  similar 
conditions  of  nearly  twice  those  indicated  in  the  table. 

In  a  slide-valve  engine  22"X28"  at  118  r.p.m.  with  a  boiler 
pressure  of  no  Ibs.  absolute,  clearance  at  8  per  cent,  a  mean 
effective  pressure  of  43.1  Ibs.  is  obtained.  By  calculation  from 
the  indicator  diagram  the  steam  consumption  was  0.64703  Ibs. 
per  stroke.  The  same  size  of  Corliss  engine  with  a  j  cut-off  and 
3  per  cent  clearance  used  0.32635  Ibs.  of  steam  per  stroke.  Each 
engine  making  14,160  strokes  per  hour,  the  steam  consumption 
will  be,  in  the  two  cases,  respectively  33.5  and  17.08  Ibs.  per  horse- 
power per  hour.  The  Corliss  therefore  saves  nearly  one-half  the 
steam  consumption  of  the  slide-valve  engine.  A  zoo-H.P.  Corliss 
will  annually  save  150  tons  of  coal  over  a  slide-valve  engine. 

Setting  the  Engine. — A  foundation  is  prepared  for  the  engine 
before  its  arrival.  Two  accurately  constructed  templates,  ordered 
from  the  manufacturers,  are  laid  in  proper  position  with  bolts 
connecting  them  long  enough  to  reach  from  the  bottom  to  the  top 
of  the  proposed  masonry  foundation.  Upon  the  lower  one  a 
foundation  of  concrete^  or  brick  is  built,  without  disturbing  the 
bolts,  to  the  level  of  the  bed-plate,  which  is  finished  and  adjusted 
ready  for  the  engine.  The  upper  template  may  be  removed  and 
the  engine  at  once  dropped  into  position  without  delay.  Tim- 
ber foundations  are  used  only  for  second-motion  haulage-engines 


n8 


MANUAL  OF  MINING. 


THE  MEAN  EFFECTIVE  PRESSURES  AND  TERMINAL  PRESSURES  AND  STEAM  CON- 
SUMPTION IN  CORLISS  NON-CONDENSING  AND  CONDENSING  ENGINES. 


%  Cut-off. 

\i  Cut-off. 

Initial 

M.E.P. 

Termi- 

Water Con- 

M.E.P. 

Termi- 

Water Con- 

r^res- 

nal 

sumption. 

nal 

sumption. 

sure. 

Pres- 

Pres 

N.  C. 

C. 

sure. 

N.  C. 

C. 

N.  C. 

C. 

sure. 

N.  C. 

C. 

60  

24.24 

36.24 

16.08 

27.6 

20.5 

29.56 

41.56 

19.77 

27-3 

21.2 

65  .... 

26.93 

38.93 

17.15 

26.5 

20.3 

32-61 

44.61 

21.09 

26.4 

21.0 

70.  ... 

29.63 

41.63 

18.23 

25-7 

20.1 

35-67 

47.67 

22.41 

25-8 

20.8 

75---- 

32-32 

44-32 

I9-3I 

24.0 

19.9 

38-72 

50.72 

23-73 

25.2 

20.6 

80.... 

35-02 

47-02 

20-39 

24-3 

19.8 

41.78 

53.78 

25-05 

24-8 

20.4 

80.... 

37-73 

49.71 

21.46 

23-7 

10.7 

44.83 

56-83 

26.37 

24-3 

20.3 

90.... 

40.41 

52.41 

22-54 

23.4 

19-5 

47-89 

59.89 

27.69 

24.0 

20.3 

95  

43.10 

55-io 

23.62 

23.0 

19.4 

50.94 

62.94 

29.01 

23.6 

20.2 

too  

45.80 

57-8o 

24.70 

22.6 

19.4 

54.00 

66.00 

30-33 

23.1 

20.1 

%o  Cut-off. 

s"Hoo  Cut-off. 

60... 

34-21 

46.24 

23-37 

27.4 

22.0 

38.18 

50.68 

27.04 

28.3 

22.8 

65.... 

37-6: 

49.61 

24.94 

26.7 

21.9 

41.80 

53-80 

28.85 

27-5 

22.6 

70  

40.98 

52.98 

26.  51 

26.1 

21.7 

45-42 

57-42 

30.66 

26.8 

22-5 

£::;: 

44-35 
47.72 

59-72 

28.07 
29.64 

25.6 
25-1 

21-5 

21.3 

49  •  °5 
52.68 

61  .05 
64.68 

32-47 
34-28 

26.2 
25-7 

22.4 
22-3 

85  .... 

51.09 

49.71 

31.20 

24.6 

21.2 

56-31 

68.31 

36.09 

25-3 

22.2 

90  — 

54-46 

66.46 

32-77 

24.2 

21.0 

59-94 

71.94 

37-90 

25.0 

22.2 

95.... 

57-83 

69.83 

34-33 

23-9 

20-9 

63-57 

75-57 

39-71 

24-7 

22.  0 

100.... 

61.20 

73.20 

35-90 

23-7 

20.8 

67.20 

79.20 

4L52 

24-5 

21.9 

tto  Cut-off. 

Yz  Cut-off. 

60.... 

41.68 

53-68 

30.66 

28.8 

23-6 

47-50 

56.50 

37.98 

29.4 

25.2 

65.... 

45-54 

57-54 

32-71 

28.2 

23.4 

51-75 

63-75 

40-52 

28.8 

25.0 

70.  .. 

49.41 

61.41 

34-77 

27.6 

23-2 

56.00 

68.00 

43-07 

28.4 

24.8 

75  

53-27 

65-27 

36.82 

27.1 

23.0 

60.25 

72.25 

45.61 

27.9 

24.7 

80.... 

57-14 

69.14 

38.88 

26.7 

22.9 

64.50 

76.50 

48.16 

27.6 

24-6 

85.... 

61.00 

73-oo 

40-93 

26.1 

22.8 

68.75 

80.75 

50.70 

27.2 

24-5 

90  

64.87 

76.87 

42-99 

25.8 

22.6 

73-oo 

85.00 

53-25 

27.0 

24.4 

95  

68.73 

80.73 

45-04 

25-5 

22.5 

77-25 

89.25 

55-79 

26.8 

24.2 

100  

72.60 

84.60 

47.10 

25.2 

22.4 

81.50 

93-5° 

58.34 

26.6 

24.1 

These,  reduced  to  coal  consumption  at  the  rate  of  10  Ibs.  of  steam  per  pound 
of  coal,  can  be  at  once  referred  to  the  coal-pile. 

The  Rotary  Effort. — The  "horse-power"  of  the  engine  as- 
sumes a  uniform  effort  of  the  engine  against  the  resistance,  but 
owing  to  varying  angles  occupied  by  the  crank-arm,  which  receives 
the  thrust  from  the  connecting-rod,  the  rotary  effort  upon  a 


POWER  GENERATION.  IIQ 

crank-pin  varies  from  one  dead-center  to  a  maximum  and  thence 
diminishes  to  a  minimum  during  each  half  of  the  revolution  or 
each  stroke.  Owing  to  this  variation  in  effort  the  cylinders 
are  usually  duplicated  side  by  side,  their  pistons  operating  upon 
their  respective  cranks  90°  apart.  This  enables  the  piston  of 
one  cylinder  at  the  middle  of  its  stroke  to  assist  the  piston  start- 
ing its  stroke. 

The  governors,  discussed  above,  check  variation  of  speed 
from  stroke  to  stroke  but  not  during  each  single  stroke.  The 
variations  in  the  rate  of  rotary  effort  upon  the  crank-pin  must  be 
kept  within  reasonable  limits  by  use  of  a  fly-wheel.  There  is 
a  varying  pressure  on  the  piston  due  to  expansion,  a  varying 
thrust  upon  the  connecting-rod  due  to  its  angularity,  and  a 
varying  tangential  effort  upon  the  crank-pin  due  to  the  angular 
position  of  the  crank-arm.  These  variations  cause  a  fluctuation 
in  the  speed  of  the  crank-pin,  the  crank-shaft,  and  of  the  hoister- 
drum  which  is  attached  to  it.  The  shorter  the  connecting-rod 
the  more  nearly  equal  are  the  forces  throughout  the  stroke  and 
in  each  half -re  volution.  A  diagram  drawn  to  represent  the 
tangential  effort  upon  the  crank-pin  by  vertical  lines  and  the 
rectilinear  motion  of  the  crank-pin  by  the  horizontal  distances; 
the  curve  shows  the  varying  efforts,  or  torque,  of  the  engine. 
If  diagrams  be  prepared  representing  the  rotary  effort  of  each 
piston  of  a  duplex  engine  and  be  superposed  upon  one  another, 
at  the  same  time  advanced  90°  to  represent  the  difference  in 
the  angle  of  the  crank-arms,  the  sum  of  the  ordinates  at  any 
one  point  represents  the  aggregate  rotary  effort  of  two  pistons 
upon  the  crank-shaft. 

An  engine  must  be  capable  of  starting  from  any  position  of 
its  piston,  or  of  its  corresponding  crank-pin,  and  have  at  that 
instant  a  rotary  effort  in  excess  of  the  resistance  on  the  crank- 
arm. 

The  Steam-turbine  may  be  employed  for  purposes  of  a  motor 
engine  where  a  high  rate  of  revolution  is  desired,  as,  for  example, 
with  a  centrifugal  pump  or  an  electric  generator.  For  these  pur- 
poses it  is  particularly  adapted. 


120  MANUAL  OF  MINING. 

The  steam-turbine  is  a  wheel  fitted  at  its  circumference  with 
numerous  blades  receiving  the  impact  of  a  jet  of  steam  flowing 
against  them  at  a  high  velocity.  The  turbines  may  be  classified 
according  to  the  path  of  the  steam,  which  may  be  radial,  outward 
or  inward,  parallel  flow  or  spiral  flow.  The  turbine  utilizing 
the  steam  which  flows  in  a  direction  at  a  constant  distance  from 
the  axis  of  the  wheel  is  the  one  most  highly  developed  at  the  pres- 
ent time.  Of  these  the  De  Laval  single  wheel  and  the  Parsons, 
Curtis,  and  Rateau  multiple  wheels  are  the  types. 

The  De  Laval  Wheel. — In  this  a  jet  of  steam  is  driven  through 
a  nozzle,  D  (Fig.  48),  where  it  expands  to  the  atmospheric  pres- 
sure, thus  creating  a  high  velocity  of  flow  whose  energy  is  imparted 
to  the  blades  of  the  wheel  producing  a  peripheral  speed  of  30,000 
feet  per  minute.  This  wheel,  being  8"  or  more  in  diameter, 
makes  15,000  or  more  revolutions  per  minute.  This  rate  being 
too  high  for  motors,  a  reduction  gear  is  employed.  Pumps  and 
electric  generators  are  attached  to  the  shaft,  making  1000  or 
more  revolutions.  In  Fig.  48  are  shown  also  details  of  its 
construction. 

The  Parsons  Steam-turbine. — This  is  a  series  of  rotary  discs 
fitted  with  blades  around  their  circumference  (Fig.  49).  Alter- 
nating with  the  discs  are  rings  of  fixed  blades  projected  inward 
from  the  casing.  The  steam  is  discharged  from  a  nozzle,  thence 
to  the  fixed  guide-blade,  which  deflects  it  to  the  rotary  blade. 
This,  after  absorbing  some  of  the  power,  again  discharges  the  fluid 
to  the  next  guide-blade,  etc.  Thirty  or  more  such  passages  are 
encountered  before  exhaust,  and  each  one  of  the  16,000  rotary 
blades  receives  a  small  amount  of  pressure,  the  aggregate  of 
which  produces  the  effective  rotation  of  the  shaft.  The  discs 
are  fitted  on  the  one  axle  and  revolve  together,  making  400  revolu- 
tions per  minute. 

The  Curtis  and  Rateau  types  of  wheels  are  very  similar,  but 
occupy  intermediate  positions  between  those  already  mentioned, 
as  to  the  method  of  utilizing  the  energy  of  the  steam. 

The  Advantages  of  the  Steam-turbine. — It  is  a  rotary  machine, 
requiring  very  small  floor  space  and  little  or  no  foundation.  It 


POWER  GENERATION. 


121 


is  much  lighter  in  weight  than  is  a  steam-engine  delivering  the 
same  amount  of  power.  These  steam-turbines  are  employed  at 
pressures  not  to  exceed  150  Ibs.  gauge.  The  steam  should  be  dry 
and  to  a  moderate  degree  is  superheated  before  being  delivered 
to  the  turbine.  Condensers  may  be  added  to  increase  their 
power. 

Its  steam  consumption  is  about  the  same  as  that  of  a  steam- 
engine    of   equal   power.     There   being  no   internal   lubrication, 


FIG.  48.— De  Laval  Wheel  Deta'ls. 

the  steam  does  not  come  into  contact  with  a  running  part  and 
its  condensed  exhaust  can  be  returned  to  the  boiler  free  from  oil. 
This  is  a  merit  of  great  importance  where  the  supply  of  feed- 
wrater  is  chemically  bad.  There  are  no  complicated  valves 
to  give  occasion  for  injury,  and  no  pistons  to  leak.  All  bear- 
ings are  easily  lubricated.  The  only  \vearing  parts  are  the 
nozzles  and  the  blades.  Turbines  can  be  had  in  power  from 
those  of  10  horse-power  up  to  one  of  7400  I.H.P. 

In  point  of  efficiency  and  commercial  utility  it  has  the  field, 
which  will  soon  be  undisputed.  Its  thermodynamic  efficiency 
is  equalled  only  by  the  gas-  or  oil-engine. 


123 


MANUAL  OF  MINING. 


v 


POWER   GENERATION.  123 

Oil-engines. — When  the  distances  are  too  remote  for  econom- 
ical admission  of  steam  or  compressed  air  through  the  pipes  or 
of  electricity  by  wires  to  pumps  and  fans,  independent  oil-  or  gas- 
engines  are  employed.  These  are  self-contained  and  are  placed 
on  the  same  foundation  with  the  machine  to  be  driven,  and  may 
even  be  portable  on  a  truck.  These  engines  require  little  atten- 
tion and  have  a  high  efficiency  in  addition  to  the  advantage  of 
portability.  The  motor  fluid  is  gasoline,  naphtha,  or  alcohol, 
which  is  held  in  a  supply-tank  from  which,  by  means  of  a  small 
pump  attached  to  the  engine,  it  is  conducted  to  the  working 
cylinder  after  having  been  atomized  or  vaporized.  It  is  then 
mixed  with  air  in  proportions  regulated  by  a  suitable  device. 
This  vaporous  mixture  is  drawn  into  the  cylinder  during  the 
first  stroke  of  the  piston,  compressed  into  a  large  clearance  space 
in  the  second  stroke,  at  the  end  of  which  the  mixture  is  ignited 
and  expanded,  the  explosive  force  of  which  drives  the  piston 
forward  on  its  third  and  outward  stroke.  On  the  return  or 
fourth  stroke  the  products  of  combustion  are  expelled  into  the 
atmosphere.  The  operation  is  then  repeated,  there  being  but 
one  working  stroke  out  of  each  four  strokes  or  every  two  revolu- 
tions of  the  fly-wheel?  This  constitutes  what  is  known  as  the 
four-cycle  engine  (Fig.  50).  Two  or  more  such  cylinders  may 
be  attached  on  the  same  shaft,  with  their  cranks  at  various  angles 
to  produce  uniform  rotary  effort  A  very  heavy  fly-wheel  is 
essential  to  carry  the  engine  over  the  three  non-working  strokes. 
It  is  started  by  hand  and  thereafter  requires  little  attention.  Its 
thermodynamic  efficiency  exceeds  that  of  the  steam-engine,  and 
it  is  useful  where  the  power  must  be  continuous,  as  for  electric 
generation  or  pumps.  Fig.  147  illustrates  one  of  these  internal- 
combustion  engines  of  the  two-cycle  type,  the  compression  being 
conducted  outside  of  the  cylinder. 

It  is  essential  that  the  amount  of  air  be  chemically  perfect  for 
complete  combustion,  when  the  initial  pressure  attained  will 
be  200  or  300  Ibs.  per  square  inch.  The  temperature  of  the 
explosion  is  over  1000°  F.  The  cylinder  is  cooled  by  the  water 
circulating  in  a  jacket,  and  the  exhaust- gases  are  delivered  into 


124 


MANUAL  OF  MINING. 


the  stack.  The  former  absorbs  from  40  to  50  per  cent  of  the 
heat  developed  in  the  combustion,  and  the  exhaust  carries  off  30 
to  40  per  cent  more,  leaving  about  20  per  cent  for  effective  work. 
The  oil-consumption  per  horse-power  hour  is  from  £  to  £  of  a 
gallon. 

Ignition  may  be  attained  by  an  electric  spark  or  a  tube  kept 
red-hot  by  a  flame  during  the  entire  running.    The  efficiency  of 


FIG.  50.— The  Four  Cycles  of  a  Gas-  or  Oil-engine. 

the  oil-engine  increases  with  the  increase  of  the  degree  of  com- 
pression at  the  end  of  the  second  stroke.  It  follows  therefore 
that  the  Diesel  motor,  which  raises  the  compression  to  700  Ibs. 
per  square  inch — nearly  the  point  of  ignition  of  the  gases — gives 
the  highest  results  (Fig.  147).  One  I.H.P.  is  obtained  from  the 
consumption  of  0.3  Ib.  oil. 

The  regulation  of  these  internal- combustion  engines  is  obtained 
by  a  governor,  which  produces  a  missfire  when  the  engine  is  under- 
loaded. 


POWER   GENERATION.  125 


REFERENCES, 

Boiler  Erection,  Adjustment,  Am.  Mir.,  July  23,  1897,  115;  Water  Tubu- 
lar Boilers,  Superiority,  Am.  Mfr.,  Jan.  23,  1897,  113;  Root  Safety  Boiler,  with 
Tables  of  Conversions,  Catalogue;  Boiler-room  Practice,  W.  T.  Edwards, 
Am.  Elect'n,  Feb.  1904;  Corrosion  of  Boilers,  H.  Ost,  Engng.,  Oct.  10,  1902; 
Boiler  Scale  and  the  Transmission  of  Heat,  Locomotive,  Sept.  1002;  Water- 
tube  Boilers,  Sci.  Am.  Sup.,  Dec.  27,  1902,  Jan.  3,  1903,  and  William  Fyvie, 
Aust.  Min.  Strand.,  Oct.  i,  1903;  Revue  Technique,  Feb.  25,  1900. 

Selecting  a  Boiler  Feed,  W.  H.  Wakeman,  Mod.  Mach.,  Mar.  1900; 
Comparative  Methods  of  Heating  Feed-water,  Power,  Nov.  1902;  Feed- 
water  Purification,  Am.  Elect'n,  Mar.  1900. 

Coal  Storage,  Waldron  Fawcett,  Sci.  Am.,  Oct.  18,  1902;  The  Valuation 
of  Coals,  E.  L.  Rhead,  Ir.  &  Coal  Trds.  Rev.,  April  3,  1003;  Coals:  Their 
Sources  and  Heating  Values,  John  W.  Langley,  Engrs.  U.  S.  A.,  April  i, 
1903;  Spontaneous  Ignition  of  Coal,  Mines  &  Min.,  Dec.  1899;  Coal  Con- 
sumption in  Central  Stations,  Alfred  S.  Giles,  Mech.  Engr.,  Jan.  23,  1904; 
Smokeless  Firing,  Colliery  Manager,  Jan.  17,  1896,  24. 

Mechanical  Stokers,  Edwin  Titts,  Pro.  Engrs.  Soc.  of  W.  Penn.,  April 
1903;  Comparative  Economy  of  Stoking  and  Hand  Firing,  Julius  Geldard, 
Mech.  Eng.,  Jan.  3,  1903;  Mechanical  Stokers  Depending  upon  the  Coking 
Method  of  Firing,  Albert  A.  Gary,  Engr.  U.  S.  A.,  April  i,  1903;  Surface 
Condensers,  C.  G.  Robbins,  Marine  Engng.,  Oct.  1902;  The  Cost  of  Steam 
Raising,  John  Holliday,  Elec.  Engr.,  Lond.,  Nov.  24,  1899;  Mechanical 
Draft,  J.  I.  Lyle,  Engr.  U.  S.  A.,  Feb.  15,  1904. 

Notes  on  Machinery  Constituting  a  Mining  Plant,  Alfred  C.  Cards,  Can. 
Min.  Rev.,  Sept.  30,  1902;  The  Development  of  Power-driven  Machinery 
in  the  Mine,  E.  H.  Roberton,  Eng.  Mag.,  Mar.  1904;  The  Designing  of  Steam- 
power  Plants,  Walter  C.  Kerr,  St.  Ry.  Jour.,  Oct.  4,  1902. 

The  Curtis  Steam  Turbine,  Engng.,  Feb.  5,  1904;  The  Steam  Turbine, 
William  Chilton,  Prac.  Engr.,  Feb.  12,  1904;  The  Turbine  Problem,  H.  F. 
Schmidt,  Am.  Elect'n,  Feb.  1904;  The  Curtis  Steam  Turbine,  W.  L.  R.  Emmet, 
St.  Ry.  Jour.,  April  n,  1903;  A  New  Steam  Turbine,  Elec.  Rev.  N.  Y., 
April  n,  1903;  The  Rateau  Turbine,  St.  Ry.  Jour.,  April  18,  1003;  The 
Steam  Turbine:  Its  Commercial  Aspect,  Edwin  H.  Sniffin,  St.  Ry.  Rev., 
Oct.  ii,  1902;  The  Westinghouse  Steam  Turbine,  Edwin  Yawger,  Am.  Mfr., 
Oct.  23,  1902;  Problems  in  Hauling  and  Hoisting,  A.  Brown,  A.  I.  M.  E., 
Vol.  XXXI,  265;  and  a  serial  on  Steam-turbine  Design,  The  Engineer,  1904. 

Slide  Valves,  Haswell,  731;  Automatic  Valves  Controlling  Engines,  Coll. 
Eng.,  April  1896,  199;  Indicator  Diagrams,  Calculation  of,  Coll.  Eng.,  Feb. 
1897,  320;  Steam-engine  Breakdowns,  Coll.  Guard.,  Nov.  13,  1896,  930; 
Shafts,  Cylinders,  Valves,  Coll.  Guard.,  Nov.  20,  1896,  965;  Tables,  etc., 


126  MANUAL  OF  MINING. 

Coll.  Guard.,  Nov.  20,  1896,  065;  Setting  Corliss  Valves,  Engineer,  Chicago, 
June  15,  1903;  Setting  Corliss  Valves,  The  Engineer,  June  15,  1903;  Setting 
Corliss  Valves,  The  Engineer,  June  i,  1904. 

Tables  of  Flow  of  Steam  and  Drop  in  Pressure  through  Pipes,  The  Engi- 
neer, Oct.  15,  1902,  697;  Governors  and  Governing  Mechanism,  Herbert 
Reed  Hall,  Prac.  Eng.,  Oct.  6,  1902;  Hints  on  the  Construction  of  Shaft 
Governors,  C.  B.  Risley,  Power,  April  1903. 

Petroleum  Engines,  Priestman,  Colo.  S.  of  M.  Quart.,  I,  1893,  65;  Cur- 
rent Practice  in  Gas-engine  Design,  H.  Lee  Koenig,  and  G.  W.  Rice,  Sib. 
Jour,  of  Engng.,  June  1903;  The  Balancing  of  Locomotive  Engines,  Engr., 
Lond.,  Feb.  28,  1902. 

Mechanical  Engineering  of  Collieries,  Serial,  J.  Futers,  Coll.  Guard.,  VoL 
LXXXVLLI,  114;  "Power  Plants"  and  Details,  A.  Meyer. 


CHAPTER  V. 

HOISTING  MACHINERY. 

Hoisting  Machinery. — The  various  mechanisms  used  for 
hoisting  minerals  in  the  mines  are  the  windlass,  the  whim,  the 
whip,  and  the  engine.  The  first-named  is  employed  for  moder- 
ate depths  of  mine  and  in  auxiliary  underground  openings  where 
neither  the  weight  to  be  hoisted  nor  its  distance  is  very  great. 
The  whim  and  whip  usually  represent  the  next  stage  in  the  de- 
velopment of  a  metalliferous  mine,  when  the  limit  of  the  former 
hand  methods  has  been  exceeded.  These  are  driven  by  a  horse, 
or  team  of  horses,  and  are  employed  for  depths  not  exceeding 
200  feet.  The  steam-engine  is  the  motor  for  all  mines  having 
a  large  output  or  a  great  depth  of  shaft,  and  even  for  a  depth 
of  but  100  feet  is  still  more  economical  than  the  simpler  methods 
mentioned  above. 

Windlasses  are  employed  for  shafts  of  moderate  depth  and  for 
winzes  underground  where  manual  labor  is  depended  upon  for 
hoisting  purposes.  The  amount  that  can  be  raised  by  two  men 
100  feet  in  eight  hours  cannot  exceed  4  tons,  allowing  for  delays, 
etc.  The  windlass  is  a  barrel,  the  wooden  cylinder,  6  inches 
to  10  inches  in  diameter,  long  enough  to  reach  across  the  shaft, 
raising  by  its  axle  on  uprights,  and  operated  at  each  end  by  cranks 
15  inches  long  set  at  right  angles  to  each  other.  Winches  and 
crabs  or  windlasses,  in  iron,  can  also  be  had  in  every  possible 
combination,  but  the  simpler  the  machine  the  less  is  the  friction, 
and  the  more  acceptable  it  is.  Each  additional  gearing  in- 
volves a  large  percentage  of  loss,  and  there  is  little  to  spare  from 
the  average  man-power  of  5300  ft.-lbs.  per  minute.  Windlasses 

127 


128  MANUAL  OF  MINING. 

are,  however,  useful  for  incidental  purposes  in  handling  heavy 
pieces  of  timber,  machinery,  and  pump-pipes.  A  given  power  can 
only  equal  a  certain  product  of  weight  and  velocity;  for  an  in- 
crease of  speed  the  weight  must  be  proportionately  diminished. 
With  a  15-inch  crank-arm,  12  revolutions  per  minute,  coil  25  feet 
of  rope  on  an  8-inch  barrel,  and  with  this  speed  of  hoist  the  greatest 
load  that  may  be  moved  under  the  circumstances  by  an  average 
laborer  is  214  Ibs.  Friction  and  stiffness  of  the  rope  will  reduce 
this.  A  i5o-lb.  load  can  be  raised  at  a  speed  of  only  35  feet  per 
minute. 

Balancing  the  Loads. — The  lengthened  windlass  barrel  re- 
ceives several  coils  of  rope,  from  the  ends  of  which  two  buckets 
are  suspended,  the  one  falling  to  balance  that  rising.  The  length 
of  rope  is  a  few  coils  only  greater  than  the  depth  of  the  shaft. 
At  the  start  the  weight  to  be  hoisted  is  only  the  contents  of  the 
tub  plus  the  rope.  This  weight  diminishes  in  rising,  till  at  the 
top  it  is  the  contents  minus  the  rope.  This  does  not,  how- 
ever, obviate  the  great  stress  from  inertia  which  arises  at 
the  moment  of  starting.  For  this  reason  single  or  double  conical 
barrels  are  used,  on  which  rope  is  coiled  in  such  manner  that 
the  empty  bucket  is  hung  from  the  larger  diameter,  while  the 
rope  from  the  loaded  tub  at  the  bottom  is  wrapped  around  the 
smaller  end.  The  tubs  balance  each  other,  but  the  empty  acts 
with  a  greater  leverage  than  the  loaded  tub,  and  thus  assists 
the  power  in  overcoming  the  inertia.  After  the  hoisting  is  under 
way,  the  empty  and  its  lengthening  rope  uncoils  with  diminish- 
ing leverage,  while  the  load  with  its  shortening  rope  gradually 
winds  on  a  larger  diameter  of  the  cone  The  buckets  do  not 
meet  in  the  middle  of  the  shaft,  but  in  the  middle  of  the  number 
of  revolutions  of  the  barrel.  In  any  event  the  pitch  of  the  cone 
must  be  calculated  for  the  given  conditions  of  depth  and  Toad, 
otherwise  its  advantage  is  manifest  only  at  the  start.  When 
properly  constructed  the  conical  drum  is  to  be  recommended. 
Not  many  attempts  are  made  to  apply  this  mechanism  to  hand 
hoisting,  but  it  is  rapidly  coming  into  favor  with  horse-  and  engine- 
power,  notwithstanding  that  it  is  dearer  than  cylindrical  drums. 


HOISTING  MACHINERY  129 

Manual  labor  is  manifestly  too  expensive  to  be  regarded  as 
any  but  a  temporary  expedient  for  hoisting. 

Whims. — When  the  height  of  he  hoist  exceeds  60  feet  or  the 
output  5  tons  per  shift,  horse-power  is  employed  to  advantage. 
The  average  horse  develops  an  effort  of  135  Ibs.  when  walking  at  a 
speed  of  180  feet  per  minute.  This  is  much  below  the  theoretical 
value  fixed  for  a  horse-power,  yet  it  represents  the  results  of  many 
tests  upon  the  energy  of  the  animal,  which  will  raise  9  tons  150 
feet  per  day. 

For  a  slightly  greater  quantity  two  horses  are  used.  But 
when  a  larger  quantity  is  to  be  handled,  a  more  efficient  and  eco- 
nomic power  is  employed.  The  utilization  of  horse-power  is 
but  an  intermediate  step  in  the  history  of  many  mines.  Where 
water  is  scarce,  fuel  dear,  and  the  transportation  of  machinery 
difficult,  the  horse- whim  serves  temporary  ends  as  a  simple  and 
tolerably  satisfactory  hoister. 

The  invariab'e  arrangement  is  a  wheel-and-axle  machine, 
consisting  of  a  drum  and  driving-beam  to  which  the  horse  is 
harnessed.  Two  sticks  6"X6"  and  9  feet  long  are  mortised  to- 
gether at  right  angles  to  each  other,  with  four  4-inch  planks 
trimmed  to  the  quadrant  of  a  circle.  These  are  held  a  foot  or 
two  apart  by  studs,  and  to  them  3-inch  plank  staves  are  spiked 
to  form  the  barrel;  which,  though  upheld  on  the  axle,  turns 
freely  about  it.  The  axle  is  a  round  2-inch  rod  stepped  in  z. 
stone  or  iron  block  and  held  at  the  top  in  an  iron  socket  on  the 
span-beam.  The  latter  is  10  inches  square,  36  feet  long,  sup- 
ported on  legs  mortised  to  it  and  braced  A  square  iron  axle- 
rod  fastened  to  the  drum  and  turning  wi:h  it  is  often  seen,  but 
is  not  so  good  as  the  free  axle.  The  entire  frame  can  be  built 
for  $ico.  The  drum  may  be  above  or  below  the  driving-beam 
(Fig.  SO- 

A  derrick  frame  is  necessary  over  the  shaft  at  a  height  suffi- 
cient for  convenience  of  handling  the  hoisted  tub  The  hoist- 
rope  passes  over  a  sheave  at  the  top  and  to  the  drum  if  set  up 
high  over  the  driving-pole,  or  under  another  pulley  at  its  base  if 
the  drum  is  close  to  the  ground.  The  latter  arrangement  is 


130 


MANUAL  OF  MINING, 


FIG.  51.— A  Whim. 


HOISTING  MACHINERY.  131 

cheaper  to  build,  but  is  wasteful  o"  power,  For  lowering  the 
tub  (Fig.  51),  a  lever  is  in  reach  of  the  driver,  by  which  the  driv- 
ing-beam may  be  disengaged  from  the  drum  and  the  tub  lowered 
by  its  own  weight,  uncoiling  the  rope  from  the  drum.  A  band- 
brake  3"  X  \"  regulates  the  speed.  The  brake  must  be  set  so  as 
to  work  with  the  motion,  not  opposite  to  it;  and  the  brake  force 
exerted  to  produce  larger  tension  in  the  driving,  not  in  the  slack, 
portion  of  the  band.  The  length  of  the  driving-beam  and  the 
diameter  of  the  drum  may  be  altered  at  will  but  the  ratio  be- 
tween them  is  also  the  ratio  of  the  speed  of  the  horse  to  that  of 
the  hoist 

If  circumstances  permit,  two  ropes  may  be  operated  from  the 
same  drum,  one  ascending  and  the  other  descending  Conical 
drums  may  also  be  employed.  Iron-framed  whims  are  on  sale, 
in  which  a  drum  is  horizontal  and  turned  by  a  bevel-gear  on 
its  axle,  fitting  to  another  at  the  central  end  of  the  drive-beam. 
While  convenient  and  easy  to  erect,  the  introduction  of  the  bevel- 
gear  involves  additional  friction. 

The  plane  of  the  derrick-pulley  should  be  tangent  to  the 
drum,  and  the  latter  far  enough  away  that  the  rope  may  coil 
and  unwind  freely  without  charing  on  its  adjoining  coil.  This 
is  accomplished  by  arranging  the  point  of  departure  of  the  rope 
from  the  pulley  at  the  same  height  as  the  central  coil.  Where 
the  full  and  empty  tubs  are  simultaneously  operated,  this  can 
be  attained  only  approximately.  If  greater  nicety  is  desired, 
the  pulley  slides  on  an  inclined  p^ne  as  the  coiling  or  uncoiling 
proceeds,  or  else  the  drum  shifts  its  position  by  turning  on  a 
screw-thread.  No  lateral  motion  is  allowed  to  the  sheave,  since 
the  rope  must  occupy  a  central  position  in  its  own  hoisting  com- 
partment. 

Tte  Hoisting-engine. — A  hoister  is  a  simple  engine  attached 
to  a  drum  on  which  a  rope  is  wound  The  combinations  are 
numerous,  from  that  of  the  upright  boiler  and  engine  on  the 
same  base,  to  the  vertical  condensing,  compound  engine  con- 
r.ected  with  drums  on  separate  foundations.  Generally  the 
cylinders  are  horizontal.  The  choice  between  horizontal  and 


MANUAL  OF  MINING. 


HOISTING  MACHINERY.  I3J 

upright  engines  is  chiefly  one  of  space  The  horizontal  engine 
is  the  cheaper,  the  s.mpler,  the  easier  to  inspect,  and  the  easier 
to  repair.  Outside  of  the  advantage  of  requiring  less  space, 
the  upright  engine  has  the  advantage  of  less  wear  on  the  cylin- 
der, and  a  more  direct  strain  upon  the  foundations.  The  types 
are  few,  from  the  slow  haulage-engines  for  tail-rope  to  the 
high-speed  hoists  with  large  drums  Most  hoisters  are  oper- 
ated intermittently;  only  in  a  few  instances  are  they  continuous. 
The  engine  must  be  simple,  safe,  under  complete  control,  capable 
of  quickly  attaining  full  speed  and  being  accurately  stopped. 

The  selection  of  the  type  will  largely  depend  upon  the  speed 
desired  for  hoisting  and  the  choice  of  valves,  condensers,  ratio  of 
expansion,  compound  or  simple  engine,  and  the  comparative  cost 
and  maintenance.  Frequently  the  question  of  economy  receives 
the  last  consideration,  simplicity  of  construction  and  safety  of 
operation  being  of  prime  importance. 

First-  and  Second-motion  Hoisters.— The  communication  of 
motion  from  piston  to  the  drum  is  direct  or  secondary  and  may 
or  may  not  be  provided  with  a  reversing  motion.  In  the  former 
class  the  piston  is  coupled  directly  to  the  drum  through  the  con- 
necting-rod. The  load  is  hoisted,  held,  or  lowered  by  steam  at 
the  will  of  the  engineer.  High  speed  is  possible  with  direct- 
acting  engine  (Figs.  52,  57  and  58.) 

Second-motion  Engines.-— All  engines  transmitting  the  power- 
io  the  drum  through  one  or  two  pinion  wheels  or  a  friction-clutch 
are  second-motion  engines  These  engines  are  slower  but  under 
greater  control  than  the  direct- acting  engines,  besides  offering 
less  risk  to  overwinding.  The  pinion-wheels  may  be  toothed 
wheels  gearing  with  others  on  the  drum- shaft,  or  friction- wheels 
bearing  against  those  on  the  drum-shaft  (Fig.  53). 

The  ratio  of  this  gear  varies  between  4  to  i  and  7  to  i,  the 
latter  ratio  being  for  very  slow  speed.  Some  quarry  engines 
are  provided  with  a  double  set  of  gear-wheels  for  two  speeds 

(Fig-  54). 

The  geared  engine  is  intended  for  slow  hoisting,  shallow 
depths,  and  small  outputs.  The  direct-acting  is  employed  for 


134 


MANUAL  OF  MINING. 


.  HOISTING   MACHINERY. 


J35 


m 


136  MANUAL  OF  MINING. 

large  outputs  and  with  a  hoisting  speed  of  never  less  than  500 
feet  per  minute.  The  limit  to  the  size  of  the  gear-wheels  is 
determined  by  the  peripheral  speed  of  the  teeth,  which  should 
not  exceed  1000  feet  per  minute. 

Second-motion  engines  may  be  operated  with  a  short  cut-off, 
which  is  not  always  possible  with  the  direct-acting  tyye.  More. 
over,  their  repairs  are  greater  and  the  operation  easier.  A  properly 
designed  second-motion  engine  may  be  employed  economically 
on  steep  slopes,  and  occasionally  for  short  shafts  where  the 
output  is  large.  A  pair  of  16"  X3o"  cylinders  on  a  three-foot 
drum  had  an  output  of  1413  tons  from  a  shaft  370  feet  deep 
in  six  hours'  time. 

Internal  Gear  Connection. — Frequently  drums  are  fitted  with 
pinion  gearing  into  an  internal  rack  on  the  drum.  This  can  be 
slipped  out  at  will,  leaving  the  free  drum  to  pay  out  the  rope  until 
the  proper  depth  is  reached,  when  the  pinion  is  returned  to  place, 
and  the  drum  is  ready  for  hoisting  from  that  depth.  This  arrange- 
ment is  in  ended  for  the  variable  lengths  of  the  hoist. 

Friction-engines  have  a  pinion-wheel  on  the  driving-shaft 
of  papier  roache"  bearing  against  an  iron  surface  on  the  drum- 
shaft.  The  surfaces  are  kept  dry  and  free  from  water,  steam,  or 
other  substances  which  reduce  the  coefficient  of  friction  and 
thus  require  an  excessive  pressure  to  move  the  requisite  load. 
Friction-engines  deliver  the  power  to  the  drum  more  quietly 
than  toothed-gear  or  direct-connected  engines,  and  therefore  pro- 
duce less  shock  upon  the  rope.  Sometimes  the  engine  is  in 
continual  rotation  with  its  pinion-shaft,  and  hoisting  is  effected  by 
bringing  the  drum  with  its  shaft  and  gearing  surface  against 
the  pinion.  For  lowering,  the  drum  is  released  from  contact 
with  its  pinion  and  revolves  freely,  being  controlled  only  by  the 
brake  manipulated  by  the  engineer.  The  risk  of  slipping  and 
the  inability  of  the  engineer  to  apply  the  requisite  pressure  makes 
the  friction-engine  unsafe. 

V  Friction-wheels. — The  surfaces  of  the  drum-band  and  the 
driving  pinion  are  grooved  circumferentially  to  give  a  tight  grip 
between  them  without  excessive  pressure  on  the  bearings.  These 


HOISTING  MACHINERY. 


138  MANUAL  OF  MINING. 

wheels  are  of  cast  iron.  The  irregular  wear  of  the  wedges,  how- 
ever, makes  them  unreliable  in  time. 

Friction-clutches.— These  engage  the  drum  externally  or 
internally.  They  are  quick,  noiseless,  and  give  little  shock. 
A  number  of  blocks  attached  to  a  band  furnish  the  frictional 
surface,  and  are  directly  attached  to  the  revolving  shaft  and  in 
continual  rotation  about  the  drum.  They  engage  the  outside 
of  the  drum  when  forced  by  a  two-armed  driver  operated  by 
bell-crank  levers  and  revolve  the  drum  (Fig.  56).  A  grooved 
sleeve  or  cone  moving  along  the  shaft  may  also  cause  a  shorten- 
ing of  the  arms  which  bring  the  band  down  on  the  drum  and 
impart  their  motion  to  it.  For  lowering,  the  band  is  released  and 
the  drum  revolves  while  held  in  check  by  the  brake. 

One  form  of  internal  friction  consists  of  driving-bands  with  four 
arms  keyed  to  the  drum-shaft  and  in  continuous  motion.  The 
drum  is  free  on  this  shaft  and  will  hoist  as  soon  as  the  band  is 
expanded  to  contact  with  its  rim,  by  sliding  the  sleeve  upon 
its  shaft  and  engaging  its  arms  against  its  inside  rim  (Fig.  55). 

The  Brake. — This  is  either  an  iron  band  with  blocks  of  hard 
wood  on  end,  or  a  V-grooved  wheel  bearing  on  a  corresponding 
surface  on  the  drum.  The  first  is  less  troublesome  and  safer.  It 
is  suitably  applied  by  a  simple  lever,  by  a  hand- wheel  and 
worm-screw,  or  an  auxiliary  steam-piston,  according  to  the  size 
of  the  hoist.  Some  large  engines  require  a  small  regulating- 
engine  to  stop  and  start  them. 

The  Unbalanced  Load  on  the  Engine. — The  resistance  which  an 
engine  must  overcome  will  include  the  unbalanced  weights,  the 
friction  and  the  bending  stress  of  the  rope.  These  having  been 
determined,  the  size  and  the  weight  of  the  rope  become  known. 

The  maximum  speed  of  hoist  is  limited  by  the  equipment 
of  the  shaft,  and  the  minimum  by  the  hourly  -output  desired. 
Owing  to  the  diminishing  weight  of  the  rope  as  the  hoisting  pro- 
gresses and  the  unequal  work  thrown  upon  the  engine,  a  constant 
speed  of  hoisting  cannot  be  maintained.  The  engine,  from  the 
beginning,  speeds  up  with  the  reducing  load,  while  the  accelera- 
tion increases  almost  to  the  end  of  the  hoist. 


HOISTING  MACHINERY. 


140 


MANUAL  OF  MINING. 


HOISTING   MACHINERY. 


141 


142  -  MANUAL  OF  MINING. 

Balancing  the  Dead  Load. — The  work  of  the  engine  divided 
between  the  live  load  and  the  dead  load  is  unequal  because 
of  the  diminishing  influence  of  the  rope  during  the  hoist.  This 
inequality  can  be  partially  reduced  by  balancing  the  weight  of 
the  rope.  The  cages  and  the  cars  carrying  the  load  may  be 
balanced  by  attaching  to  another  rope  on  the  same  drum  and 
engine  a  similar  cage  and  car.  The  weights  of  the  two,  going 
in  opposite  directions  balance  one  another,  though  the  weight  of 
the  rope,  its  friction,  and  its  bending  stress  are  still  unbalanced; 
so  also,  the  weight  of  the  mineral  being  hoisted. 

The  balancing  of  cage  and  car  may  be  accomplished  by  the 
use  of  two  cylindrical  drums  (Fig.  57),  a  conical  or  a  fusee  drum 
or  a  counterbalancing  system. 

Balancing  by  Conical  Drum. — The  method  of  balancing  by 
the  use  of  a  cone  or  a  fusee  depends  on  the  change  of  leverage 
of  the  two  loads.  The  load  is  hoisted  from  the  small  diam- 
eter' of  the  drum,  while  the  descending  rope  is  suspended  from 
the  larger  diameter  of  the  cone.  As  the  hoisting  proceeds,  the 
ascending  rope  winds  upon  an  increasing  diameter,  while  the 
descending  rope  and  cage  arc  gradually  operating  from  a  shorter 
radius  of  the  drum.  Equalization  is  somewhat  maintained  and 
an  easy  start  is  thus  made. 

A  p'ain  conical  drum,  to  be  effective  as  a  counterbalance, 
requires  an  angle  so  great  as  to  be  dangerous  on  account  of  the 
liability  of  he  rope  to  slip.  Any  advantage  to  be  derived  from 
having  a  safe  angle  being  trifling,  it  is  really  not  worth  the  risk. 

The  cylindrical  drum  admits  of  extension  of  the  hoisting 
depth  but  the  conical  or  fusee  drum  is  built  for  a  specific  condi- 
tion of  load,  speed  and  depth. 

The  Conical  Drum. — This  is  used  when  the  rope  is  heavy 
and  the  economy  of  accurate  counterbalance  is  clearly  indicated 
as  practicable  and  will  warrant  the  increased  cost.  It  is  much 
larger  than  ihe  cylindrical  drum,  even  if  its  mean  diameter  is 
the  same.  The  depth  of  shaft  or  the  point  of  hoist  must  be 
fixed  to  maintain  equalization  with  a  given  cone  or  fusee.  Hoist- 
ing from  several  landings  would  not  be  economical  without  the 


HOISTING   MACHINERY. 


J43 


use  of  additional  counterpoises.    The  conical  drum  is  employed 
single  or  double  (Figs.  53,  58,  and  59). 

In  designing  the  single  or  double  conical  drum  the  hori- 
zontal distance,  P,  between  the  centres  of  the  two  consecutive 
grooves  or  coils,  and  the  vertical  distance,  p,  between  them  must 
be  in  such  ratio  that  the  inclination  shall  not  exceed  30°;  p+P 
must  be  less  than  0.577.  The  minimum  diameter,  2r,  of  the 
smaller  end  is  fixed  by  the  rope,  and  the  larger  diameter,  2d,  is 
dependent  upon  it. 

Let  c  =  the  circumference  of  rope  circle  =  3. 1416(27+ #); 
if  =  0.2830^; 
L  =  length  of  the  drum  between  the  first  and  last  grooves. 

The  lenglh  can  be  determined  when  the  number  of  grooves 
has  been  calculated  from  the  conditions  given. 


FIG.  59.— A  Double  Conical  Hoist. 

The  Fusee. — This  drum  is  designed  to  give  an  equalization 
of  the  dead  load,  including  the  rope,  at  all  points  throughout  the 
journey  of  the  cage.  It  may  be  determined  by  the  use  of  the 
following  formula: 


I44  MANUAL  OF  MINING. 

After  one  revolution  (RD+A  and  the  weight  of  one  turn  of  rope 
at  the  small  end)  p  must  balance  (4  + the  weight  of  one  turn  of 
rope  at  the  large  end)  p'  and  still  equal  Ad  =  M.  For  each 
revolution  throughout  the  journey  this  balance  must  be  main- 
tained. This  is  satisfied  by  the  following  formula,  in  which 
Z  is  the  total  arc  of  revolution  described  by  a  point  on  the  rope 
at  the  small  end  from  the  beginning  to  the  end  of  the  hoist: 


R  is  the  weight  of  one  foot  of  rope. 

The  curve  of  the  drum  is  then  constructed  by  substituting 
different  values  between  r  and  d,  which  are  the  limits  of  p,  placing 
the  second  member  equal  to  27tn,  and  solving  for  n,  the  number 
of  grooves.  The  various  assumed  va'ues  for  p  are  radii  of  the 
curve  at  the  various  points  along  the  axis  which  are  at  a  distance 
from  the  initial  point  equal  to  the  horizontal  pitch,  P,  multiplied 
by  n.  The  curve  so  plotted  is  the  section  of  the  drum  which 
will  fulfil  the  conditions,  furnishing  an  equalization  throughout 
the  journey. 

Let  D  =  2ooo  feet,  R  =  $  Ibs.,  A  =4000  Ibs.,  r=$>  inches; 
then  M  =  40,000  ft.-lbs.,  and  d  becomes  10  feet.  Solving  the 
equation  and  equaling  with  2rw,  ^  =  55.8  revolutions  between  the 
initial  and  final  points  of  winding.  Where  d  =  g.  n  becomes 
53.3  revolutions  from  the  start;  where  d  =  8,  ^  =  49.5:  where  J=7, 
n=44-4;  and  where  d=$,  here  have  been  35.1  coils  of  rope. 

The  Reel. — This  is  a  barrel  with  spider  arms  on  either  side, 
between  which  a  flat  steel  rope  is  wound  in  consecutive  layers. 
It  may  be  so  devised  that  the  moment  of  resistance  will  be  nearly 
constant  The  thickness  of  the  coiling  rope  may  be  so  selected 
that  it  will  increase  the  leverage  of  the  load  at  the  same  rate  as 
the  weight  of  the  load  decreases,  in  which  event  the  work  of 
hoisting  is  uniform.  If  however,  uniform  hoist'ne;  cannot  be 
obtained  throughout  the  trip  with  a  reasonable  thickness  of  rope, 


HOISTING  MACHINERY. 


'45 


FIG.  60.— Plan  of  a  Geared  Reel  Hoister,  showing  the  Brake  Connections. 


146  MANUAL  OF  MINING. 

the  design  must  be  carefully  considered  and  the  engine  tested 
after  being  placed  (Fig   60). 

The  smallest  allowable  diameter  of  barrel  is  fixed  by  the  rope, 
and  the  number  of  layers,  n,  of  the  rope  is  equal  to  (d—r)  +  t. 
The  diameter  of  the  coil,  2d,  when  the  cage  is  at  the  top  of  he 
hoist  may  be  ascertained  by  the  use  of  the  formula  P/  =  3.i4i6 
(d1—  r2).  Given  /  (f"  to  £"),  the  value  for  d  may  be  obtained; 
or,  having  decided  upon  d,  t  can  be  ascertained. 

Counterbalance. — The  balancing  of  the  rope  weight  and  its 
resistance  is  accomplished  either  by  the  varying  weight  of  an 
attached  counterpoise  to  the  cages  or  drum,  or  by  varying  the 
leverage  on  the  hoisting-shaft  of  a  fixed  weight.  The  Koepe 
and  Whiting  systems  and  the  chain  balance  are  of  the  first  type, 
and  that  installed  at  the  Camphausen  mine  is  of  the  second  type. 
This  makes  the  moment  of  the  static  load  during  the  hoist  con- 
stant and  equal  to  that  of  weight  of  mineral  The  best  results 
are  obtained  from  a  counterbalance  when  the  hoist  is  always 
from  one  and  the  same  level 

The  Koepe  System  of  Winding. — This  form  of  counterpoise 
consists  in  extending  a  tail  rope  from  the  bottom  of  one  cage, 
under  a  sheave  at  the  bottom  of  the  shaft,  up  to  the  floor  of  the 
other  cage.  The  usual  hoist-drum  is  replaced  by  a  simple  grooved 
sheave  Fig.  61.  Not  only  is  the  weight  of  rope,  cage,  and  car 
in  each  compartment  constant  and  the  work  of  the  engine  nearly 
uniform,  but  the  comparative  lightness  of  the  sheaves  and  the 
absence  of  heavy  coils  revolving  with  the  drum  give  a  correspond- 
ingly less  inertia,  and  hence  require  less  engine  power,  than  with 
the  ordinary  systems.  The  net  load  which  can  be  carried  on  the 
rising  cage  cannot  exceed  the  friction  caused  by  the  aggregate 
weights  suspended  from  the  upper  sheave,  or  else  it  will  slip. 
A  positive  grip  must  be  taken  on  the  rope  by  the  driving  mechan- 
ism to  prevent  its  slip  or  creep  on  the  sheaves.  Winding  the 
main  rope  on  a  pair  of  cylindrical  drums,  or  wrapping  it  a  few 
times  over  a  pair  of  multiple-grooved  sheaves,  as  is  the  European 
practice  in  power  transmission,  Chapter  IX,  would  accomplish 
this.  Should  such  a  slip  take  place,  the  indicators  would  fail 
to  give  the  correct  location  of  the  cage. 


HOISTING  MACHINERY. 


147 


This  system  meets  almost  all  the  requirements  of  a  perfect 
equilibrium,  is  highly  efficient,  insures  against  overwinding,  de- 
creases wear,  and  disposes  of  an  enormously  heavy  drum  by  using 
a  sheave  instead.  But  the  lower  sheave  blockades  the  bottom 
of  the  shaft.  The  tail-rope  may  be  a  discarded  hoisting-rope, 


FIG.  6 1. —The  Koepe  System  of  Winding. 

for  it  has  no  work  to  do  beyond  supporting  its  own  weight.    The 
method  is  of  limited  application  as  to  depth. 

In  a  shaft  1260  feet  deep  a  saving  of  n  per  cent  in  power  is 
effected.  With  sheaves  8£  feet  in  diameter  and  8-inch  axle, 
main  rope  175  Ibs.  per  foot  and  tail-rope  1.55  Ibs.,  cage  and  car 
2100  Ibs.,  load  2750  Ibs.,  depth  of  shaft  2500  feet,  and  velocity 


148 


MANUAL  OF  MINING 


of  hoist  2500  feet  per  minute,  250  horse-power  are  saved  over 
that  required  to  operate  the  same  loads  without  this  balance. 
The  Whiting  system  (Fig.  62),  a  development  of  the  Koepe 


FIG.  62.— The  Whiting  System  of  Balance. 

system,  uses  two  sheaves  and  dispenses  with  the  heavy  coils  of 
rope  revolving  with  the  drum. 

Chain  Counterbalance. — Another  form  of  counterpoise  is  a 
heavy  chain  wound  on  a  secondary  drum  on  the  main  shaft  by 
a  cable-rope  (B,  Fig.  63).  The  chain  hangs  in  an  auxiliary 
shaft,  or  a  ladderway  in  the  hoist-shaft,  and  is  equal  in  weight 
to  the  full  depth  of  hoist-rope  A.  As  the  loaded  cage  begins  its 
FIG.  63.  FIG.  64. 


Counterbalancing  the  Rope  by  Chain. 

hoist  the  entire  weight  and  length  of  chain  is  hanging  from  its 
drum  to  assist  the  engines  in  balancing  the  rising  load.  With 
the  continuance  of  hoist  the  cable  B  pays  out  chain,  which  coils 


HOISTING  MACHINERY.  149 

on  the  bottom  of  the  auxiliary  shaft.  As  the  two  cages  approach 
one  another  the  hoist-ropes  nearly  balance,  the  chain  is  coiling 
at  the  bottom,  and  the  cable-rope  is  being  paid  out.  When  the 
cages  are  midway,  the  entire  chain  is  coiled  there  and  all  the 
cable  is  also  paid  out.  The  counterbalancing  cable-rope  is  n^w 
wound  under  the  drum  and  lifts  the  chain,  whose  weight  now 
balances  the  descending  rope  to  the  end  of  its  journey.  At  all 
points  in  the  hoist  the  amount  of  suspended  chain  balances  the 
differences  between  the  weights  of  the  two  pendent  hoist-ropes  A. 

In  a  given  case  the  chain  counterbalances  a  load  of  3500  Ibs.,. 
:age  1300  Ibs.,  car  1540  Ibs.,  in  a  shaft  3400  feet  deep;  the  radius, 
of  the  hoist-drum  is  6  feet,  and  that  of  the  chain-drum  27  inches. 
At  the  beginning  of  the  hoist  8  tons  of  chain  are  suspended  from 
the  auxiliary  drum.  After  400  feet  of  hoist,  12,134  Ibs.  are  sus- 
pended; at  400  feet  further,  8400  Ibs.;  at  2200  feet  below  the 
surface,  4666  Ibs.;  at  1800  feet  down,  933  Ibs.  are  yet  uncoiled; 
and  at  the  end  of  the  hoist  the  full  weight  of  the  chain  is  again 
suspended.  The  bore-hole  down  which  the  chain  hangs  is  638 
feet  deep;  51  H.P.  are  required  to  overcome  the  friction  due  to  this 
increased  weight,  but  345  H.P.  are  saved  by  balancing  the  rope. 
In  another  shaft,  2200  feet  deep,  a  counterbalance-rope  700  feet 
long  and  a  chain  580  feet,  weighing  4320  Ibs.,  are  employ edj 
the  size  of  the  chain  is  graduated  to  meet  the  varying  weights, 
and  for  104  feet  is  of  f-inch  links,  for  162  feet  of  f -inch,  and  for 
314  feet  of  f-inch. 

In  Despre's  method  an  endless  cable  has  the  balancing  chain, 
and  is  given  one  or  two  turns  on  the  drum  between  the  full  and 
the  empty  rope  (Fig.  64).  It  passes  half-way  down  an  auxiliary 
shaft.  One-quarter  the  way  down  the  chain  is  swung  on  a 
pivot,  and  fixed  to  the  cable  so  that  the  other  end  of  the  chain  is 
at  the  top  of  the  auxiliary  shaft  at  the  beginning  and  end  of  the 
hoist. 

The  Camphausen  System. — A  counterpoise  (W,  Fig.  65)  is 
suspended  from  a  supplementary  spiral  drum  mounted  on  the 
engine-shaft.  The  cable-ends  are  attached  to  the  small  and  large 
ends  of  the  spiral.  It  winds  itself  on  the  descending  side  simul- 


I5o  MANUAL  OF  MINING. 

taneously  with  the  winding  on  the  ascending  side,  and  thus  main- 
tains a  balance  by  the  differential  lever-arm  of  its  weight.  Until 
the  cages  meet  in  the  hoistway,  the  lower  branch  of  the  cable 
travels  faster  than  the  upper.  During  the  balance  of  the  hoist  the 
upper  winds  faster  than  the  lower.  For  a  depth  of  1900  feet 
the  auxiliary  shaft  was  250  feet.  The  installation  is  inexpen- 
sive and  economical. 


FIG.  65. — The  Camphausen  System  of  Counterbalancing. 

A  very  extended  discussion  of  the  operations  of  hoisting  is 
to  be  found  in  a  monograph  of  the  Institute  of  Mining  and 
Metallurgy,  XI,  Part  I. 

Designing  Roisters. — This  step  consists  in  calculating  the 
dimensions  of  the  two  cylinders,  which  are  capable  (i)  of  starting 
the  load  from  the  weakest  position  of  their  combined  crank- 
pins,  and  (2)  of  maintaining  an  average  speed  of  hoist  with  it 
commensurate  with  the  desired  output.  The  mine  output  is 
controlled  by  the  capacity  of  the  shaft  or  the  haulage-way,  which 
in  turn  depends  on  the  size  of  the  car  and  the  number  of  cars 
delivered  per  hour. 


HOISTING  MACHINERY.  151 

Data  Necessary  in  Designing. — From  a  shaft  or  a  steep 
slope  the  cars  arrive  at  the  surface  singly;  from  a  gentle  slope 
the  cars  come  in  trains.  The  dimensions  of  the  carriers  (car, 
bucket,  or  skip)  and  their  capacity  are  known  or  assumed.  The 
load,  q,  which  a  single  car  can  carry  is  also  known,  hence  the 
load  which  falls  upon  the  hoisting-  or  haulage- engine  can  be 
determined.  This  may  be  divided  into  a  dead  load  and  a  live 
load,  the  former  being  the  cage,  car,  skip,  or  bucket,  while  the 
live  load  is  the  mineral  and  the  friction  resistances. 

The  speed  depends  upon  the  character  of  timbering.  A 
mine  employing  buckets  has  usually  a  poorly  timbered  shaft,  in 
which  the  speed  cannot  exceed  300  feet  a  minute;  with  skips  or 
slope  carriages  the  timbering  must  be  of  a  better  character,  and 
the  maximum  rate  of  1000  feet  is  allowed;  if  cages  are  em- 
ployed, the  shaft  is  well  timbered  and  the  speed  may  reach  3000 
feet  per  minute. 

The  depth  of  the  shaft  is  known;  hence  the  time  required  per 
trip  is  definitely  fixed;  and  if  to  it  be  added  the  time  lost  at  the 
top  and  at  the  bottom  while  the  buckets,  cars,  or  carriages  are 
being  exchanged  and  loaded,  the  time  occupied  in  each  round 
trip  in  each  hoistway  will  be  known.  The  time  lost  in  filling  an 
attached  bucket  at  the  bottom  or  dumping  it  into  the  top  is  from 
3  to  5  minutes.  If  the  buckets  are  detached  and  promptly  replaced 
by  loaded  buckets,  the  time  is  much  less.  Skips  loaded  from 
chutes  are  almost  as  time-consuming  as  buckets.  The  skip 
occupies  2  to  3  minutes  to  side-track.  A  car  may  be  run  on  or 
taken  off  the  cage  or  a  slope  carriage  in  25  seconds.  These 
delays  seriously  reduce  the  capacity  of  the  shaft  and  the  possible 
output.  Hence  for  large  mines  cages  and  cars  are  resorted  to. 
The  influence  of  this  delay  on  the  aggregate  output  is  diminished 
by  increasing  the  number  of  compartments  in  the  shaft  or  by 
increasing  the  velocity.  The  former  is  preferable. 

Formulae  for  Determining  the  Number  and  Duration  of 
Trips. — By  the  following  equations  may  be  determined  the  time 
occupied  per  round  trip  in  each  hoistway  and  the  number  of 
trips  possible  per  hour  in  each  hoister: 


I52  MANUAL  OF  MINING. 

Let   2  =the  number  of  hoisting- compartments; 
D  =the  depth  of  the  shaft  in  feet; 
v  =  velocity  of  the  hoist,  feet  per  minute; 
i/  =  velocity  of  lowering  per  minute; 
Q  =output,  tons,  per  hour; 
0=load,  tons,  per  trip; 
/=the  minutes  to  load  and  unload  at  the  top  or  the 

bottom ; 

T  =4z=minutes  per  round  trip  in  any  hoistway; 
A  =the  interval  in  minutes  between  the  arrival  of  each 

cage,  skip,  or  bucket  at  the  shaft  top; 
w=number  of  trips  per  hour  in  each  hoistway; 

n=Q  +  zq,    A~     and     T^~+^+zt. 

Thus  in  a  poorly  timbered  shaft,  if  only  one  bucket  be  run  up  and 
down  without  detaching,  the  output  from  a  3oo-foot  shaft  is  about 
3.6  tons  per  hour.  With  three  buckets  in  constant  use,  each 
holding  600  Ibs.,  the  hourly  product  cannot  exceed  6  tons.  With 
excellent  timbering  double  the  speed  may  be  permitted,  in  which 
event  three  buckets  will  deliver  7.2  tons  per  hour  at  the  surface. 
So  becomes  evident  the  influence  which  the  loss  of  time  at  land- 
ings has  on  the  output  when  even  doubling  the  speed  increases 
the  former  only  one-fifth.  An  additional  hoistway  were  better. 

EXAMPLE. — For  an  output  of  360  tons  in  ten  hours  a  rate  of  360  feet  is 
necessary  with  but  one  hoistway;  with  two  hoistways  and  cars  holding  3000 
Ibs.  each  the  rate  need  be  only  144  feet  per  minute,  because  of  the  time  saved 
at  the  changes.  2=1;  /=o.4i;  D=3oofeet;  <7=i.5;  £=36;  ^=24;  ^  =  250 
min.;  7=2.5  min-'>  whence  ^=^,  =  366  feet.  With  Z=2,  A  becomes  2.5 
and  «=i2.  T=5  min.;  whence  7>=7/,=  i44  feet.  The  horse-power  in 
the  two  cases  would  be  100  and  32  respectively.  This,  at  the  coal -pile,  amounts 
to  40  tons  coal  per  month. 

Determining  the  Size  of  Rope.— The  rope  supports  the  total 
load,  the  friction,  and  the  starting  stresses.  The  latter  is  larger 
than  is  usually  suspected.  A  gentle  start  gives  a  stress  of  2  per 
cent  only,  whereas  a  start  from  12  inches  of  slack  rope  would 
make  the  stress  1.4  times  as  great  as  the  actual  load.  The  stress 


HOISTING  MACHINERY.  1 53 

at  starting  is  usually  assumed  at  10  per  cent  of  the  gross  load, 
including  friction.  The  total  tension  produced  by  these  resist- 
ances must  not  exceed  one- third  of  the  ultimate  strength  of  the 
rope. 

The  Bending  Stress  on  the  Rope. — The  rope  must  be  sufficient 
for  the  maximum  stress  falling  on  the  rope  after  hoisting  begins, 
including  the  stress  induced  by  bending  the  rope  over  its  sheave 
or  drum.  This  sum  is  not  to  be  over  one-third  of  the  ultimate 
strength  of  the  rope.  Frequently,  with  a  small  drum,  the  bend- 
ing stress  may  far  exceed  that  due  to  the  load.  The  value  for 
the  bending  stress  is  represented  by  the  formulae  in  Chapter  VII. 

The  Available  Strength  of  a  Rope. — After  deducting  for 
bending  stresses,  this  should  be  at  least  one-eighth  of  its  ultimate 
strength.  The  size  required  may  be  ascertained  from  manu- 
facturers' tables. 

EXAMPLE. — A  6-strand,  ig-wire  rope  of  cast  steel  i  inch  diameter  will  be 
required  if  it  is  to  support  a  load  of  7270  Ibs.  over  a  sheave  44  inches  diameter. 
k— 14,606,  load  and  friction =8000,  and  maximum  load  is  22,606.  Ultimate 
strength  must  be  67,800  Ibs.  Had  the  drum  been  84  inches  diameter,  k 
would  have  been  5583  Ibs.  and  have  required  an  ultimate  strength  40,752  for  a 
J-inch  rope. 

Determining  the  Dimensions  of  Cylinders. — The  size  of  the 
rope,  the  dimensions  of  the  drum,  and  the  values  for  the  various 
loads  and  resistances  to  be  overcome  having  been  determined, 
the  area  and  stroke  of  the  piston  is  made  sufficient  to  start  the 
load  with  the  allowable  initial  pressure  and  hoist  it  at  the  desired 
speed.  The  work  of  the  engine  must  also  be  calculated  for  the 
average  resistances  at  the  mean  velocity  per  minute.  The  cylinders 
should  not  be  too  large,  because  the  throttling  that  must  ensue 
for  the  economic  work  is  wasteful  of  steam,  for  the  condensation 
would  be  excessive.  Neither  can  they  be  calculated  for  an  over- 
load, as  then  they  might  not  meet  the  probable  emergency.  The 
cylinders  should  be  apportioned  largely  according  to  the  conser- 
vative judgment  of  the  engineer.  Care,  however,  must  be  taken 
to  assure  a  sufficiently  large  starting  moment  of  the  engine  to  pick 
up  the  load.  The  mechanical  efficiency,  m,  may  be  used  as  0.7 


154  .MANUAL  OF  MIXING. 

for  a  slide-valve  engine  and  a  cut-off  of  one-half.  The  Corliss 
engine  may  be  calculated  to  have  an  efficiency  of  0.85  with  a  cut- 
off of  one-third,  for  a  normal  load. 

The  Hoisting  Capacity  of  an  Engine. — The  following  equa- 
tions will  then  be  employed  for  determining  the  depth  to  which 
a  given  engine  can  work  within  the  normal  limits  of  boiler  pres- 
sure which  is  available  and  the  average  r.p.m.  for  that  class 
of  engine.  They  will  also  serve  to  calculate  the  required  dimen- 
sions of  one  or  two  cylinders  for  given  conditions: 

Z,  =  the  length  of  the  drum,  inches; 
r=the  smaller  diameter  of  the  drum,  inches; 
</=the  larger  diameter  of  the  drum,  inches; 
w  =  the  number  of  grooves  in  the  drum; 
X  =  the  diameter  of  the  rope,  inches; 
R  =  the  weight  of  the  rope  in  pounds  per  foot. 

v=the  velocity  of  hoist,  feet  per  minute,  — '• — - — ; 

6 

A  =the  weight  of  bucket,  cage,  car,  skip,  or  carriage  Ibs.; 
M  =  unbalanced  load  on  the  drum  at  the  start; 

K = the  moment  of  resistance  at  starting; 
TF=the  average  load  during  the  hoist; 

F= the  friction  of  the  gross  load,  including  the  acceleration 
of  the  moving  parts; 

£=gear  ratio,  second-motion  engine,  =  -; 

y= number  of  teeth  on  the  gear-wheel; 
x = number  of  teeth  on  engine-shaft  pinion. 

The  capacity  of  the  engine  may  be  known  from  its  indicator 
cards  or  from  a  knowledge  of  the  cylinder  dimensions  and  the 
proposed  boiler  pressure,  P,  and  engine  speed.  The  velocity 
for  a  given  output,  number  of  trips,  and  depth  will  have  been 
determined.  Then 

H =o. 0000039666$  k2mpN = Wv. 


HOISTING  MACHINERY.  155 

Or,  the  value  for  k  and  s  may  be  ascertained  for  a  given  set  of 
conditions  of  load  and  speed  as  well  as  for  P  and  N.  The  ratio  of 
s  to  k  is  assumed  between  1.3  and  2,  according  to  the  class  of 
engine  desired. 

With  a  direct-acting  engine  g=i.    In  all   cases  the  mean 

sVg 
piston  speed  is  0.31841—1 — . 

Hoisting  on  Inclines. — If  the  hoist  is  not  in  a  vertical  shaft, 
the  load,  M ,  must  be  multiplied  by  the  sine  of  the  angle  of  the 
slope  (the  percentage  of  inclination),  and  the  value  for  F  is  multi- 
plied by  the  cosine  of  the  slope  angle,  in  calculations  for  the 
cylinder  dimensions. 

The  coefficient  of  friction,  /,  is  4  per  cent  with  a  bucket  in  a 
vertical  shaft,  6  per  cent  with  a  cage  or  skip,  and  8  per  cent  with 
a  bucket  in  a  slope.  All  of  these  are  reduced  to  the  pull  at  the 
circumference  of  the  drum. 

The  Starting  Moment  of  an  Engine. — In  designing  an  engine  to 
start  the  unbalanced  load,  M,  it  is  necessary  to  determine  by 
the  equilibrium  of  moments  the  value  for  the  moment  of  the 
load,  friction,  and  starting  stress,  and  equate  this  with  the  mini- 
mum rotary  effort  upon  the  crank-pin.  The  starting  moment 
of  the  engine  is  ascertained  by  the  substitution,  in  the  following 
formula,  of  the  value  for  c,  corresponding  to  the  given  cut-off 
and  ratio  of  connecting-rod  to  crank-arm. 

The  minimum  rotary  effort,  R,  of  a  duplex  single-expansion 
engine  =  CsPJPmg  =  R,  inch-pounds.  The  maximum  rotary  effort 
=R' =  C1'sPik2mg.  s  and  k  are  in  inches,  Px  is  the  net  initial 
pressure,  P—B,  and  C,  the  coefficient  representing  the  minimum 
rotary  effort  as  ascertained  from  the  following  table  for  the 
particular  ratio  of  connecting-rod,  l}  to  crank-arm,  a,  and  a  piston 
clearance  of  7  per  cent. 

The  values  given  are  the  rotary  efforts  exerted  upon  the 
two  crank-pins  for  each  pound  of  net  initial  pressure  for  each 
inch  of  one  piston  diameter,  the  pair  having  their  cranks  at 
quarters.  The  clearance  is  assumed  to  be  7  per  cent,  and  a  back 
pressure  of  18  Ibs.  absolute  is  allowed.  If  the  engine  is  a  con- 


'56 


MANUAL  OF  MINING. 


densing  engine,  with  3  Ibs.  absolute  back  pressure,  the  values  for 
C  and  C  are  somewhat  larger  than  those  in  the  table.  A  pair 
of  cylinders  of  10  inches  diameter  and  20  inches  stroke  with 
connecting-rods  50  inches  long,  receiving  an  initial  steam  pres- 
sure of  100  Ibs.  absolute,  will  have  a  minimum  rotary  effort 
equal  to  48,480  in.-lbs.,  if  the  cut-off  be  assumed  at  £  'and  the 
back  pressure  be  3  Ibs. 

•TABLE  or  MAXIMUM  AND  MINIMUM  ROTARY   EFFORTS  FOR   CERTAIN  RATIOS 
OF  CONNECTING-ROD  TO  CRANK-ARM.     DUPLEX  ENGINES. 

Values  for  Cl  and  C. 


Apparent 
Cut-off. 

/=4.5<J. 

l=5-5d.     ' 

;=7a. 

o.oo 

o.s;6i5 

0.3074 

0.5600 

0-3974 

0.5580 

0-3974 

\ 

0.4664 
0.4380 

0-3633 
0.2472 

0.4606 
0.4365 

0.3688 
0.2499 

0.4586 
0-4350 

0.3696 

0.2550 

\ 

0.4256 

0.1878 

0.4193 

0.1923 

0.4124 

0.1966 

0.3893 

0.1469 

0.3818 

°-I473 

0.3786 

0.1512 

The  hoisting-engine  must  be  able  to  start  the  load  from  any 
point  in  the  shaft  and  at  any  point  in  its  stroke,  and  hence  should 
have  in  its  weakest  position  a  rotary-effort  moment  greater  than 
that  of  the  maximum  load  which  may  fall  upon  it.  Doubtless 
all  the  masses  involved  will  have  attained  sufficient  velocity 
in  one  revolution  to  produce  an  average  effort  on  the  applied 
moments  from  both  cylinder-cranks.  The  engine-friction  moment 
decreases  during  the  hoist  nearly  commensurate  with  the  load 
moment. 

Hoisters  cannot  be  regulated  by  fly-wheels,  and  their  sole 
means  of  maintaining  a  uniform  rotary  effort  is  that  which  is 
afforded  by  the  drum  or  a  weight  of  the  connecting-rod.  If 
these  are  made  heavy,  the  velocity  may  be  somewhat  uniform. 
But  inasmuch  as  the  engine  must  furnish  an  excess  of  power 
at  the  beginning  which  is  50  per  cent  more  than  the  average, 
and  the  inertia  of  the  rope,  sheaves,  drum,  etc.,  must  be  over- 
come, and,  later,  the  kinetic  energy  absorbed  by  the  brake,  some 


HOISTING  MACHINERY.  157 

form  of  energy  accumulator  should  be  devised  for  the  purpose 
of  economizing  power. 

Determining  the  Size  of  the  Drum. — The  minimum  allowable 
diameter  of  the  cylindrical  drum  is  fixed  by  the  size  of  the  rope 
(Chapter  VII),  it  must  be  more  than  forty  times  the  diameter  of 
the  rope — one  hundred  times  the  diameter  would  be  preferable, 
since  thereby  it  would  reduce  the  bending  stress  of  the  rope  and 
increase  the  net  working  load.  The  smaller  end  of  the  cone 
or  fusee,  and  of  the  barrel  of  the  reel,  is  also  fixed  by  the  same 
ratio.  The  desired  hoisting  speed  also  influences  the  final  de- 
cision as  to  the  diameter.  The  drum  or  drums  are  geared  to  the 
•same  shaft  and  placed  centrally  between  the  cylinders. 

The  length  of  the  drum  is  controlled  by  the  depth  of  the  hoist 
and  its  location  with  reference  to  the  shaft.  On  cylindrical 
drums  the  rope  coils  are  contiguous  to  one  another.  Conical 
drums  are  provided  with  spiral  grooves.  Reels  have  the  suc- 
cessive coils  of  rope  winding  on  one  another.  The  length  of 
the  drum  and  its  distance  from  the  shaft  must  be  such  that  the 
acute  angle  from  the  sheave  to  the  ends  of  the  drum  must  not 
exceed  6°.  When  this  fleet  angle  exceeds  the  limit,  the  rope 
mounts  its  adjoining  coil,  and  provision  must  be  made  for  guid- 
ing the  rope  upon  sheave  and  drum,  or  the  drum  must  be  set 
back  from  the  shaft.  In  the  first  case  the  engine  may  be  set 
on  a  carriage  to  move  with  a  varying  position  of  the  rope  on 
the  drum. 

The  cases  are  rare  where  but  a  single  rope  with  a  single  cylin- 
drical drum  is  sufficient.  Two  ropeways  are  the  rule  with  two 
independent  cylindrical  drums;  or  a  double  cylindrical  drum,  two 
conical  drums,  or  a  double  cone,  the  fusee,  or  two  reels  are 
employed  according  to  the  requirements  in  the  given  case.  These 
may  be  operated  independently  of  one  another,  though  provision 
is  always  made  for  locking  the  pair  of  drums  together  to  obtain 
simultaneous  hoist  and  lowering.  This  is  true  whether  the  engine 
be  a  direct-acting  or  a  second-motion  engine. 

The  General  Formulae  for  Hoisting  Calculations. — The  follow- 
ing table  then  gives  the  value  for  the  starting  resistance  of  the 


158  MANUAL  OF  MINING. 

load  for  any  variety  of  drum,  the  moment  of  which  is  equated 
with  the  starting  effort  of  the  engine,  to  determine  the  dimensions 
of  the  cylinders.  The  value  for  the  unbalanced  load  to  be  hoisted 
throughout  the  trip  at  a  requisite  velocity  is  also  given  for  the 
selected  drum.  Then  the  resistance  M+F,  multiplied  by  the 
velocity  v,  is  equated  with  the  value  for  engine  power,  to  deter 
mine  the  engine  dimensions  for  the  dynamic  load. 


THE  STARTING  MOMENTS,  WORKING  LOADS,   ETC., 
ENGINES  AND  DRUMS. 


WITH  VARIOUS  HOISTING 


Maximum 
Unbalanced 
Load.  M. 

Friction  of 
Load,  F. 

Minimum  Admissible 
Radius  of  Drum  d, 
Inches. 

One-cylinder  drum..  . 

A+KD  +  ^ 

if/ 

if)ogDg 

***"      N 

Two-     "           "    ... 

RD  +  2oooq 

(M  +  2A)/ 

" 

Double  conical  drum. 

a 

" 

zA  +  2Oooq 

Fusee             

,000, 

(M+2A  +  RD)/ 

RD  +  A 

Reel  

« 

" 

r+nt;V3.82Dl  +  r* 

" 

(M  +  2A  +  2RDV 

25* 

Number  of 
Grooves,  n. 

Length  of 
Drum,  L, 
Inches. 

Starting  Moments, 
K.  Inch-lbs. 

One<ylinder  drum.  .  . 

1.  0098!)  m 
d 

nX 

(M+F)* 

Two-     "           "    .  .  . 

3.81960  m 

2nX 

V+p 

Double  conical  drum  . 

3.8it)6£> 

«~.) 

-A(l-t)i 

Fusee  . 

\\TF)  ~f4>(jf)  ~d* 

— 

v+'+StM 

12.566^         12.566^ 

Reel  

d~r      3.82!) 

- 

Koepe  system  .  . 

IX+IV 

HOISTING  MACHINERY.  159 

The  following  examples  will  suffice  to  illustrate  the  use  of 
the  formulas  and  tables  for  calculating  the  work  of  hoisting. 

i.  Duplex  engine,  single  drum. 

In  a  shaft  400  feet  deep  an  output  of  100  tons  per  ten  hours 
is  desired  from  a  single  compartment.  The  load  in  a  car  is 
|  of  a  ton;  weight  of  car  and  cage,  1800  Ibs.  Assuming  a  duplex 
engine,  single  drum,  and  full  steam  pressure  available  for  starting, 
required  the  size  of  the  cylinders,  number  of  trips,  etc. 

A  iQ-wire  rope  at  starting  (Chapter  VII),  on  an  8-foot  sheave, 
suffers  a  stress  K=A  +  2oooq+F+RD.  Assume  the  rope  to 
weigh  i  Ib.  per  foot.  The  total  load  (1800+1500+  400)  (i.i)  = 
4070  Ibs.  This  requires  a  f-inch  cast-steel  rope  weighing  0.9  Ib. 
per  foot,  because  the  total  load  must  be  less  than  \  the  ultimate 
strength,  which  must  be  28,490  -Ibs. 


1200. 


, 
r+o.^x 


3200  Ibs.  is  less  than  one-third  the  ultimate  strength. 

Output  per  hour,  10  tons;  <?=f;  hence,  with  single  cage,  «  =  i4 
trips;  time,  4.3  min.  for  round  trip.  Allowing  for  loss  at  top 
and  at  bottom  of  30  seconds  each,  the  hoisting  time  is  1.65  each 
way  and  ^=360  feet  per  minute.  A  drum  of  4  feet  diameter 
makes  28.6  r.p.m.  Let  /=o.i,  ^=0.75,  the  engine  stroke  =1.5^, 
and  the  mean  cut-off  =£. 

(1800+400+  1  500)  (1  +  0.1)360  =44.4  H.P.; 
0.0000039668  &  2Nmp  X2  =44.4; 

whence  kap  =174,012  and  £=13.5  inches. 

With  7  per  cent  clearance  ^=0.8925(100)—  18=71.25  Ibs.  per 
square  inch.  Cylinder  is  i7£"X26",  nominally,  with  counter- 
bore  at  each  end  of  if  inches. 

For  starting,  the  dimensions  should  be  (page  10) 

CsPlk2mg=R=g'j,'jio  inch-pounds; 

C  =0.3688;  g  =  i;  Pi=82;  whence  k*s  =4308  and  £  =  14.  25  inches. 
2.  Double  drum  and  duplex  engine. 


160  MANUAL  OF  MINING. 

A  shaft  is  2000  feet  deep;  car  and  cage,  5000  Ibs.;  load,  2\ 
tons;  velocity  of  hoist,  2000  feet;  gear  ratio,  i;  drum,  60  inches 
radius;  engine-stroke,  48  inches;  modulus,  0.8;  admission 
steam  pressure,  100  Ibs.  gauge;  back  pressure,  5  Ibs.;  cut-off 
average,  £;  sheave,  12  feet;  and  /,  o.i.  Required  the  size  of 
ip-wire  rope,  diameter  of  engine  cylinder,  and  output  of  mine. 

At  the  same  velocity  of  lowering  as  of  hoist,  the  round  trip  is 
made  in  three  minutes,  including  stops.  In  each  hoistway  20 
trips  are  made  hourly.  The  output  is  1000  tons  daily. 

The  starting  stress  is  (5000+ 5000 +.&D)(i.i)  +  .K';  assuming 
a  weight  of  7- wire  rope  of  3  Ibs.  per  foot,  and  a  diameter  of  if 
inches,  the  load  is  17,600  Ibs.,  and  the  bending  stress  13,060  Ibs, 
The  aggregate  starting  stress  must  be  less  than  one-third  the 
ultimate  strength.  That  for  a  if -inch  cast-steel  rope  is  62  tons. 

Dimensions  of  cylinders  for  normal  running: 

.F  =  (6ooo+ 15,000)0.1  =2100  Ibs.; 

(6000+5000+2100)2000  =  26,000,000  ft.-lbs.  =793.6  H.P.; 
2  Xo.ooooo3g66sk2Nmp  =  793.6. 

The  depth  of  shaft  equals  in  this  case  the  velocity  of  hoist, 
hence  the  number  of  coils  =  the  r.p.m.  • 

N  =63.6  r.p.m.;  piston  speed  =508.8;  whence  k2p=  40,966. 
For  an  average  cut-off  \  with  3  per  cent  clearance, 

p  =0.8658(100+ 14-7)-  5  =99-3; 

k2  =397  and  diameter  of  cylinder =23  inches. 

At  63.6  r.p.m.  for  ^=2000  feet,  the  diameter  of  drum  is  9.91 
feet.  The  starting  resistance  is  17,600  X59-5  =0.3688  X48  Xio9-7 
Xo.8XiX£2;  whence*  =26  inches. 

With  these  dimensions  of  cylinders  the  M.E.P.  at  the  average 
rate  may  be  56  Ibs.,  corresponding  to  a  cut-off  of  0.18  for  3  per 
cent  clearance: 

56=0(114.7)- 5,  whence  C  =0.532. 

Or,  at  a  cut-off  of  £,  an  initial  pressure  of  71  Ibs.  absolute  will 
suffice.     56=0.8658(^-5;  P  =  7i. 


HOISTING  MACHINERY.  161 

The  normal  rate  of  running  being  at  \  cut-off,,  the  Corliss 
engine  of  26"  X4&",  giving  733.3  H.P.,  will  consume  17,600  Ibs. 
steam  per  hour. 

3.  Double  conical  drum  and  duplex  engine. 
Same  specification  as  in  2. 

The  smallest  diameter  of  the  drum  is  assumed  at  80  inches,, 
for  the  rope  of  if  inches.  The  angle  of  the  cone  is  14°  30';. 
/>=o.47;  P=i.825;  c=  230.6  inches;  ^'=4.178. 

To  maintain  equal  moments  of  resistance  when  the  load  is  at 
top  and  at  bottom, 

,    5000  +  2  X6ooo  +  2  X  5000 

a  = —  • 40  =72  inches. 

2X5000+5000 

n  =  —  — 9       ^°°°=68.2   grooves;    Z=  1.825X68.2=10  ft.  4.5  in. 

This  is  rather  a  long  drum.     To  obtain  a  fleet  angle  of  6°  it 
must  be  100  feet  from  the  shaft. 

The  horse-power  required  for  the  normal  rate  of  running  is,  • 
as  before,  763.9;   ^=68.2  r.p.m.;   piston  speed,  545.6  feet;    and 
/>=99-3  Ibs.  per  square  inch;  k  becomes  19.5  inches. 

The  starting  resistance,  neglecting  the  lengths  of  rope  above 
the  surface  landing  to  the  head  sheave,  requires  a  diameter,  k> 
of  15.6  inches 

(5000+6000+  5000X1.1)40—  5000(0.9)72 

=  .3688  X48  X  109.7  Xo.8  Xk2~ 

4.  Reels  with  duplex  engine. 
Same  specifications  as  in  2. 

Barrel  diameter  is  40  inches  and  flat  rope  is  jXsJ,  weighing 
4.2  Ibs.  The  bending  stress  is 

89465* 

R+ 2.231 
the  diameter  of  the  outer  length  of  rope  is 

d=V^S~2Dl+r2=6s  inches; 
and 

n  =  2(65—  20)  =90  coils. 


162  MANUAL  OF  MINING. 

If  the  cylinders  are  direct-connected,  the  piston  speed  will 
be  720  feet  per  minute,  whence  £=17.08  inches.  The  above 
speed  is  rather  high.  If  the  barrel  be  enlarged  to  60  inches, 
then  d  becomes  97.5  inches,  AT  =60  r.p.m.,  and  &  =  2of  inches. 

The  starting  moment  of  the  engine  at  an  assumed  cut-off  of 
£  would  be 

R  =0.3688  X48  X82  X57I.8  Xo.8  X2  =1,328,000  inch-pounds. 
The  moment  of  the  starting  resistance  is 

(16,000)  (1.1)20-  (5000X0.9)65  =94,700  inch-pounds. 

5.  Koepe  system  with  duplex  engine. 

Same  specifications  as  in  2. 

The  load  upon  the  engine  is  at  all  times  5000+  the  friction  of 
the  gross  weight,  moving  at  2000  feet  per  minute.  Neglecting 
the  weight  of  the  1  2-foot  sheaves,  and  assuming  the  rather  large 
frictional  coefficient  of  o.i, 

.F=  (5000+  6000+  io,ooo)(o.i)  =2100  Ibs.; 

(5000+2100)2000 

33,000 


430.3  =2 

JV=  53.05  r.p.m.,  if  g  =  i;  and  the  piston  speed  =424.4  feet, 
whence  k  becomes  17.05  inches. 

If  the  drum  is  geared  with  a  ratio  of  2,  k  would  become  14 
inches.  Then  (5000+2700)72  =o.^6S8pk2smg,  the  minimum 
rotary  effort  of  the  engine  with  a  cut-off  of  \  being  0.3688. 

Electric  Holsters  may  be  located  in  relation  to  incline  or 
shaft  exactly  where  they  can  be  operated  to  the  greatest  advan- 
tage, considering  convenience  of  position,  without  the  pipes  insep- 
arable from  air-  or  steam-hoisters.  Their  speed  is  constant, 
the  motor  is  economical,  and  the  only  power  consumed  is 
that  actually  required  to  handle  the  load.  One  objection  to 
their  use  is  the  large  gear  ratio  which  is  required  to  bring  the 


HOISTING  MACHINERY.  163 

speed  of  motor  down  to  speed  of  bolster  (Fig.  66).  But  they 
have  none  of  the  ills  of  the  engine,  with  its  tremendous  vibration, 
rattle,  heat,  and  noise  of  exhaust,  or  of  the  compressed-air  hoist, 
which  has,  in  addition,  the  extremely  cold  exhaust. 


The  motors  for  electric  hoist  are  either  of  the  continuous- 
current,  series-  or  compound-wound  types,  or  of  the  alternating- 
current  induction  type,  designed  for  either  constant  or  variable 
speed,  and  geared  to  the  drum.  They  are  fitted  with  appropriate 
rheostats.  The  direct- current  series  motor  is  particularly  adapted 
to  cases  which  demand  a  very  high  starting  torque  and  where 
closeness  of  speed  regulation  is  not  imperative.  The  compound- 
wound  motor  should  be  employed  when  there  is  liability  of  the 
motor  racing  on  throwing  off  the  load.  It  has  a  high  starting  torque, 
and  its  speed  will  not  rise  above  a  certain  definite  value  when  the 
mechanical  load  is  thrown  off. 

The  variable- speed  induction  motor  is  best  used  for  hoisting. 
Its  starting  torque  is  quite  high  and  its  speed  regulation  sufficiently 
flexible  to  make  it  very  desirable. 


1 64  MANUAL   OF  MINING. 

Calculating  the  Size  of  Wire  for  a  Hoister. — One  of  the  three 
variable  quantities  are  to  be  determined:  the  horse-power  of  the 
motor,  the  load,  M,  and  the  speed  of  hoist,  v. 

The  mechanical  efficiency,  m,  for  the  direct-current  motors 
is  0.85  and  may  be  regarded  as  unity  for  the  induction  motors. 

The  effective  horse-power  of  the  motor  must  equal 

0.00003032;  (M + F) 
and 

M+ F  =  33000  X  horse-power  X  — . 

In  determining  the  size  of  wire  for  the  direct-current  circuit 
the  formulae  given  in  the  next  chapter  will  be  used. 

EXAMPLE. — It  is  desired  to  operate  a  30-horse-power  moto  at  a  distance  of 
500  feet  from  the  generator  by  a  direct  current  of  440  volts.  Then  D=  500;  C, 
for  i -horse-power  (Chapter  VI),  is  2  amperes  for  440-volt  circuit.  Let  V,  the 
loss  in  volts,  be  10  per  cent,  or  44  volts.  Then,  for  a  30-horse-power  motor, 
30  X  2  amperes  ( =  60  amperes)  will  be  required.  Substituting  in  the  formula,  the 
wire  must  have  14,300  circular  mils  and,  according  to  the  table,  will  be  of 
No.  8  size.  The  weight  of  the  wire  will  be  69  Ibs.  in  the  total,  and  its  cost 
$17.25. 

If  the  current  be  a  three-phase  alternating,  the  value  for  K  being  assumed 
as  17,  a  No  10  wire  with  11,600  circular  mils  will  be  sufficient.  Its  total 
length  of  1500  feet,  weighing  73  Ibs ,  will  cost  $18.25. 

REFERENCES. 

Foundations,  Lecture,  Coll.  Eng.,  April  1896,  207;  Hoisting,  Trans.  Am. 
Soc.  C.  E.,  XXIX,  695;  Mining  Machinery,  Henti  Deschamps,  Rev.  Univ. 
d.  Mines,  Jan.  1902;  Some  Notes  on  Ceryan  Underground  Hoisting  Prob- 
lems on  the  Whitwatersrand,  A.  W.  K.  Pierce,  Trans.  Am.  Inst.  Elec.  Engrs., 
Oct.  1903;  The  Mechanical  Engineering  of  Modern  Collieries,  William 
Bardill,  Coll.  Guard.,  Mar.  20,  1903;  The  Mechanical  Equipment  of  Col- 
lieries, George  W.  Winstanley,  Coll.  Guard.,  Feb.  26,  1904;  Mining  Machin- 
ery, Butte,  Montana,  Coll.  Eng.,  1896,  25. 

Compound-winding  Engine,  description  of  one  at  Cardiff,  Trans.  M.  & 
M.  Eng.,  XLV,  205;  Electrical-winding  Engine,  Dr.  R.  Hertzfeld,  Engng., 
Feb.  12,  1904;  Hoisting  Engine,  Direct-acting,  111.  Min.  Inst.,  I,  145;  Reel 
and  Hoist  at  Boston  and  Montana,  E.  &.  M.  Jour.,  Mar.  20,  1897,  285;  A 
Single-engine  Plant,  L.  Sup.  Min.  Inst.,  IV,  1896,  81;  Engine  Breakdowns, 


HOISTING  MACHINERY.  165 

Michael  Longridge,  Mech.  Engr.,  Sept.  26,  1903;  and  Trans.  A.  I.  M.  E., 
XXXI,  348. 

Equalizing  Load  on  Hoister,  A.  Rogers,  A.  I.  M.  E.,  XVII,  305;  Koepe, 
A.  I.  M.  E.,  XVII,  429;  Koepe,  Col.  Eng.,  Oct.  30,  1896,  150;  by  Balance 
Chain,  Trans.  N.  of  Eng.  M.  Inst.,  XLVI,  Part  III,  56,  with  calculations  of 
Methods,  S.  of  M.  Quart.,  1889,  259;  Counter  Balancing,  Tapering  Rope, 
S.  of  M.  Quart.,  April  1889,  260;  Air-shait  ior  Winding,  Coll.  Guard.,  Nov. 
27,  1896,  ion;  Electric  Hoist  Description,  Coll.  Guard.,  Feb.  26,  1897,  397. 

Loss  of  Load  Due  to  Motion  in  Shaits,  Coll.  Guard.,  Vol.  LXXX,  530; 
The  Dynamos  of  Winding  Engines,  ColL  Guard.,  Vol.  LXXXVI,  842. 


CHAPTER  VI. 

ELECTRIC   GENERATION  AND    WATER-POWER. 

The  Application  of  Electricity  to  Mining.  —  Though  the 
utility  and  economy  of  electricity  for  underground  work  have  long 
been  recognized  by  the  mining  engineer,  the  conservatism  forced 
upon  him  by  restrictive  legislation  has  made  its  installation  under- 
ground one  of  slow  growth.  Its  economy  and  convenience  have 
been  recognized;  its  adaptability  to  all  operations  of  mining, 
lighting,  .blasting,  coal-cutting,  hoisting,  haulage,  etc.,  its  flexibility 
to  all  distances,  is  conceded;  it  never  vitiates  the  air,  as  do  steam- 
engines,  nor  does  it  fog  or  chill  it,  as  does  compressed  air;  and 
it  is'capable  of  automatic  regulation,  requiring  but  a  small  number 
of  employees.  The  low  cost  of  maintenance  and  repairs,  its  high 
economy,  as  compared  with  other  motors,  and  the  fact  that  there 
is  no  leakage  of  power  when  the  motor  is  not  in  use,  gives  positive 
assurance  of  future  benefit,  especially  where  the  power  is  to  be 
intermittently  applied.  It  is  capable  of  transmission  over  great 
distances  with  small  loss,  and  to  localities  where  the  haulage  of 
coal  for  steam  generation  is  exceedingly  difficult.  In  this  field 
it  is  unrivalled.  The  development  of  the  steam-turbine  and 
impulse-wheels  makes  possible  direct  connections  with  the  electric 
generator,  forming  an  ideal  combination  in  point  of  economy  of 
power  and  space. 

Two  serious  objections  obtain  against  electric  installations 
for  mining  purposes,  and  they  are  all  important.  The  first  is  the 
risk  of  fire  and  shock  of  those  coming  in  contact  with  the  electric 
current.  They  cannot  be  wholly  eliminated  without  very  great 
cost  of  copper  and  its  insulation.  Nevertheless  it  is  possible  to 

166 


ELECTRIC  GENERATION  AND   WATER-POWER.  167 

insulate  to  a  high  degree,  to  place  the  wires  remote  from  acci- 
dental contact  by  the  men,  and  to  reduce  the  fire  risk  by  the  use 
of  non- sparking  devices  and  enclosed  machinery.  The  partial 
success  in  this  direction  has  removed  the  chief  objections  to 
underground  installations.  A  less  potent  objection  is  that,  elec- 
tricity not  being  applicable  for  every  operation  of  mining,  the 
owner  is  compelled  to  maintain  two  separate  power  systems 
where  percussion-drills  are  needed  and  large  bodies  of  water  are 
to  be  pumped. 

Electrical  Units. — The  electric  units  are  the  ampere  and  the 
volt,  corresponding  to  quantity  and  pressure.  The  unit  of  resist- 
ance, R,  to  the  electric  flow  is  the  ohm,  whose  value  corresponds  to 
a  unit  electromotive  force  and  permits  the  flow  of  a  unit  of  current. 
The  standard  ohm  is  the  resistance  offered  by  a  column  of  pure 
mercury  at  o°  C.,  of  uniform  cross-section,  106.3  cms.  long  and 
14.4521  grs.  weight. 

Electromotive  Force. — This  is  the  electric  pressure  which  forces 
a  current  through  a  resistance,  or  is  the  difference  of  potential 
between  its  termini.  The  unit  of  electromotive  force,  E.M.F., 
is  that  pressure  which  will  force  a  unit  current  through  a  unit 
resistance.  The  unit  is  the  volt,  which  is  0.6974  of  the  difference 
of  pressure  between  the  poles  of  a  Clarke  cell  at  15°  C.,  whose 
E.M.F.  is  1.434  volts. 

Current  Capacity. — An  electric  current,  /,  has  unit  strength 
when  a  resistance  of  one  ohm  will  afford  an  E.M.F.  of  one  volt 
between  its  ends.  The  unit  is  the  ampere,  that  current  which 
will  electrolytically  deposit  silver  at  the  rate  of  0.0011118  gr. 
per  second. 

The  quantity  of  electricity  which  passes  through  a  given 
cross-section  of  an  individual  circuit  in  /  seconds  when  a  current 
of  /  amperes  is  flowing  is  equal  to  //  units.  The  unit  is  the 
coulomb,  whose  value  equals  the  flow  of  one  ampere  for  one 
second.  One  ampere  for  one  hour  equals  3600  coulombs. 

Capacity,  C,  is  the  property  of  a  material  for  holding  a  charge 
of  electricity.  A  condenser  has  unit  capacity  when  a  constant 
of  electricity  will  charge  it  to  a  potential  of  one  volt.  The  farad 


1 68 


MANUAL   OF  MINING. 


is  the  name  of  this  unit,  but  since  its  value  is  much  greater  than 
actual  practical  values,  its  millionth  part,  the  micro-farad,  is 
used  as  the  practical  unit. 

Electric  Energy,  W,  represents  the  work  done  in  a  circuit 
or  conductor  by  a  current  flowing  through  it.  The  joule  is  the 
name  of  this  unit,  its  value  being  the  work  done  by  the  passage 
of  one  ampere  through  one  ohm  for  one  second.  Electric  power, 
P,  is  one  joule  per  second  and  represents  the  passage  of  one 
ampere  of  current  under  a  pressure  of  one  volt.  Its  value  is 
measured  in  watts.  One  watt  per  second  is  a  joule. 

i  watt  =0.7373  ft.-lbs. 
746  watts  =  i  horse-power. 

In  order  to  avoid  the  use  of  large  numbers,  the  term  kilowatt, 
or  looo  watts,  is  universally  used  in  power  and  lighting.  Allowing 
for  15  per  cent  drop,  the  following  table  has  been  computed,  show- 
ing the  quantity  of  current  necessary  to  deliver  one  horse-power 
at  the  various  stated  pressures: 

To  DEVELOP  ONE  MECHANICAL  HORSE-POWER  AT  THE  MOTOR  SHAFT. 


8  amperes 

7 

4 

3-5 

2 
I.8S 

i-75 
1.6 


no-volt  circuit. 
125 

220 
250 
440 

475 
500 

55° 


Ohm's  Law. — Resistivity,  p,  is  the  specific  resistance  of  a 
substance,  and  is  the  resistance,  in  ohms,  of  a  cubic  centimeter 
of  material  to  a  flow  of  current  between  opposite  faces.  In 
any  electrical  circuit,  the  current  which  flows  equals  the  electro- 
motive force  divided  by  the  resistance.  Expressed  in  the  form 
of  an  equation,  Ohm's  law  is  IR  =E,  in  which  7  is  the  number  of 
amperes  flowing  in  an  undivided  circuit,  E  the  algebraic  sum 
of  all  the  E.M.F.  in  the  circuit,  and  R  the  sum  of  all  the  re- 
sistances in  series  in  the  circuit. 

Since  conductors  offer  more  or  less  resistance  to  the  passage 


ELECTRIC   GENERATION   AND   WATER-POWER.  169 

of  current,  there  will  be  a  drop  or  fall  of  potential,  Ed,  along  the 
circuit,  the  amount  of  which  will  depend  upon  the  resistance  of 
the  conductor  and  the  current  that  is  flowing: 

Ed=IR. 
The  resistance  of  a  conductor  may  be  found  from  the  equation 


in  which  p  is  a  constant,  termed  the  resistivity,-whose  value  depends 
upon  the  material  and  the  temperature  of  the  conductor;  L  is 
the  length  in  centimeters,  and  A  the  cross-section  in  square  cen- 
timeters. The  resistance  of  pure  metals  increases  with  a  rise 
in  temperature.  A  rise  of  i°  C.  increases  the  resistance  by  0.004 
times  that  at  o°  C. 

Circular  Mils.  —  A  circular  unit  is  a  circle  o.ooi  inch  in  diam- 
eter. A  conductor  one  foot  long  and  one  circular  unit  cross- 
section  is  called  a  mil-foot. 

Electric  Currents.  —  Two  forms  of  currents  are  generated 
and  transmitted  for  power  purposes  —  the  continuous  current 
and  the  alternating  current.  In  the  former,  known  also  as  the 
direct  current,  the  electric  fluid  maintains  a  flow  in  one  direction 
through  the  circuit.  An  alternating  current  is  one  which  changes 
the  direction  of  its  flow  at  regular  recurring  intervals.  Dynamos 
generating  electrical  energy  are  of  two  classes,  according  to  the 
currents  to  be  produced. 

Electric  Generator.  —  The  continuous-current  dynamo  con- 
sists of  field-  magnets,  an  armature,  and  a  commutator.  The  field- 
magnets  are  poles  of  iron  arranged  in  pairs.  The  dynamo  may 
have  two  field-magnets,  when  it  is  known  as  the  bipolar  machine, 
or  several  pairs  of  field-magnets,  then  called  a  multipolar  ma- 
chine. In  the  first  case  the  armature  revolves  between  the  oppo- 
site ends  of  two  magnets,  the  other  ends  being  directly  connected. 
In  the  multipolar  machine  the  magnets  project  inward  around 
a  ring,  or  casing,  inside  of  which,  and  nearly  touching,  the  arma- 
ture revolves.  Each  pole  is  wound  with  wire,  through  which  a 


MANUAL   OF  MINING. 


current  flows  to  maintain  the  magnetism  in  the  poles.  The 
number  of  poles  may  be  2,  4,  6,  8,  or  more,  the  wire  being  wound 
right-handed  about  the  first  pole  and  left-handed  about  the  second 
in  order  to  produce  with  the  same  current  the  alternate  polarity 
in  the  respective  piece  . 


FIG.  67. — A  Drum  Armature. 

The  armature  may  be  of  the  drum  or  the  ring  type,  the  former 
being  stronger  and  generally  employed  for  power  transmission. 
A  rotary  shaft  (Fig.  67)  is  built  to  a  solid  core  of  thin  plates  of 
soft  iron  projecting  radially  and  insulated  from  one  another  by 
mica  or  paper.     They  are  bolted  together 
with  the  surface  slotted  longitudinally.      In 
the  slots  are  laid  coils  of  wire  whose  project- 
ing ends  are  connected  with  copper  strips 
composing  the  commutator.     Fig.  68  illus- 
trates the  ring  type  of  armature. 

The  commutator  has  an  equal  number 

FiG.68.-Rh^Armature.  of  c°PPer  insulated  strips,  held  firmly  in 
place  by  a  bronze  sleeve.  Brushes  collect 
the  current  from  the  commutator,  one  set  for  each  pair  of  poles. 
They  are  composed  of  wire,  or  of  small  blocks  of  carbon  coated 
wi.h  copper. 


ELECTRIC   GENERATION  AND    WATER-POWER. 


The  Winding  of  Generators. — According  to  the  method  of 
winding  the  coils,  there  are  series-wound  machines,,  shunt-wound 
machines,  and  compound-wound  machines. 

In  the  series-winding,  the  arrangement  of  coils  provides  for 
the  whole  current  flowing  continuously  and  directly  from  one 
brush  through  the  winding  of  the  magnets,  then  through  the 
external  circuit  and  back  to  the  other  brush.  The  wire  around 
the  field-magnets  is  of  large  diameter  and  low  resistance.  This 
form  of  winding  is  suitable  only  where  a  constant  current  ia 
required,  as  for  arc  lighting  (Fig.  69). 

In  the  shunt-winding,  two  paths  are  open  to  the  current. 
Leaving  one  brush,  one  branch  flows  through  the  external  circuit 
and  the  other  through  the  field-magnets.  Both  join  at  the  other 
brush  before  returning  to  the  armature.  This  also  gives  constant 
E.M.F.,  but  with  varying  conditions  in  the  external  circuit. 
The  fields  are  wound  with  a  large  number  of  turns  of  small  wire 
(Fig.  70). 


FlG.  69. — Series  Dynamo.     FlG.  70. — Shunt  Dynamo.         FIG.  71. — Compound 

Dynamo. 

The  compound-winding  is  a  combination  of  both  the  shunt 
and  the  series.  The  dynamo  gives  a  constant  E.M.F.  The  shunt 
coils  are  of  numerous  turns  of  thin  wire  overlaid  by  a  few  turns 
of  series  coils  of  thick  wire.  Both  are  coupled  up  (Fig.  71). 

The  Theory  of  the  Dynamo.— The  dynamos  are  driven  by 
belts  from  the  engine,  or  may  be  directly  coupled  to  an  engine, 
water-wheel,  or  steam-turbine.  Belt  connection  is  cheaper  than 
the  direct  connection. 

The  rotation  of  the  armature  presents  successively  the  plates. 


172  MANUAL  OF  MINING. 

of  soft  wire  in  apposition  to  the  poles  of  the  field-magnets.  This 
develops  in  each  coil  about  the  corresponding  plate  a  small 
current  of  electricity,  which  is  conveyed  to  their  ends  at  the  com- 
mutator. The  aggregation  of  the  currents  thus  successively  pro- 
duced in  the  multitude  of  coils  by  a  high  rotary  speed  is  a  cur- 
rent of  high  potential  and  large  quantity,  which  may  be  drawn 
by  the  brushes  to  the  external  wire  circuit.  The  E.M.F.  developed 
is  proportional  to  the  number  of  turns  of  wire  in  the  rotating 
armature. 

Two  wires  from  the  brushes  to  the  electric  motor  or  lamps 
complete  the  external  circuit  for  the  current  flow. 

Alternating-current  Machines. — The  wiring  of  the  alternating 
machine  is  similar  to  that  of  the  direct-current,  but  the  commutator 
is  replaced  by  two  insulated  rings  on  the  armature  shaft,  one 
end  of  the  numerous  armature  coils  being  attached  to  one  ring 
and  the  other  end  to  the  other  ring.  The  field-magnets  on  a 
bronze  base  around  the  rotor  constitute  the  stator.  Their  north 
and  south  poles  alternate,  their  windings  being  excited  by  a 
similar  continuous-current  machine  driven  from  the  alternator. 

The  armature  of  the  alternating  machine  is  called  the  rotor. 
It  carries  the  secondary  circuit  and  has  as  many  coils  systematic- 
ally arranged  around  the  laminated  soft  steel  discs  as  there  are 
pairs  of  field-magnets.  In  its  revolution  an  electric  current  is 
generated  in  the  armature  coils,  which  changes  its  direction  as 
often,  during  each  revolution,  as  there  are  pairs  of  field-magnets. 
This  gives  rise  to  the  alternating  currents. 

If  a  second  set  of  coils,  similar  to  the  first,  but  independent 
of  them,  be  wound  in  the  spaces  between  the  first  set  on  the 
armature,  then,  by  the  revolution  of  the  rotor,  are  induced  two 
distinct  sets  of  currents,  one  leading  the  other  by  one-half  a  period. 
Such  a  current  is  known  as  a  two-phase  current.  If  three  sets 
of  independent  coils  be  arranged  equally  around  the  armature, 
then  a  three-phase  current  will  be  obtained,  with  its  pulsations 
having  their  maximum  one- third  of  a  period  from  one  another. 
The  three-phase  current  is  sometimes  called  a  polyphase  or 
multiphase  current. 


ELECTRIC  GENERATION  AND   WATER-POWER.  173 

The  Frequency. — The  value  of  the  current  varies  in  the 
different  intervals  from  zero  to  a  maximum,  diminishes  with  a 
like  regularity  to  zero,  and  rises  to  the  same  maximum  in  the 
opposite  direction,  returning  finally  to  zero.  One  such  com- 
plete interval  is  a  cycle,  the  tilde  (~)  being  used  to  denote  it. 
The  time  in  which  the  cycle  of  changes  of  such  currents  is  com- 
pleted is  known  as  its  period.  The  number  of  cycles  com- 
pleted  per  second  is  its  frequency.  The  frequency  usually 
adopted  in  practice  for  power  transmission  is  usually  between 
25  and  40  cycles.  The  number  of  cycles  completed  in  a  minute 
is  known  as  its  rate  of  alternations. 

As  the  points  of  maximum  pressure  are  removed  120°  from 
that  of  the  other  circuits,  the  algebraic  sum  of  the  phase  cur- 
rents, when  balanced,  is  at  every  instant  equal  to  zero.  Two 
wires  should  be  employed  for  each  of  the  currents  and  three 
return  and  three  direct  wires  for  the  three-phase  system.  But, 
for  the  reason  stated  above,  the  three  return  wires  may  be  dis- 
pensed with.  Three  wires  then  serve  for  the  practical  three- 
phase  circuit. 

The  Double-current  Generator. — This  is  wound  to  produce 
either  or  both  varieties  of  current,  .and  is  used  where  the  cost 
of  two  separate  dynamos  has  deterred  many  from  introducing 
electric  power.  There  is  a  definite  ratio  between  the  voltages 
on  the  direct  and  alternating  sides.  For  the  two-phase  it  is 
0.76  of  that  of  the  direct  current  which  it  could  produce,  and 
0.65  of  the  voltage  if  a  three-phase  current.  Hence  the  double- 
current  dynamo  is  a  low-voltage  machine. 

The  difference  of  potential  which  can  be  obtained  from  a 
D.C.  generator  cannot  exceed  1000  volts,  because  of  the  ina- 
bility to  prevent  short-circuiting  at  the  commutator.  The  alter- 
nating-current machine  is  not  limited  in  voltage  from  this  source 
of  leakage. 

A  serious  defect  of  the  D.C.  machines  is  the  sparking  of 
their  brushes,  due  to  faulty  alignment,  an  insufficient  bearing 
or  imperfect  contact,  and  the  short-circuiting  in  the  windings. 
These  faults  suggest  simple  remedies  without  further  mention. 


MANUAL   OF  MINING. 


The  heating  of  the  coils  occurs  when  any  current  is  forced  through 
the  coils,  but  no  harmful  results  ensue  unless  the  temperature 
rises  above  80°.  When  exceeded,  the  insulation  softens  and  is 
burned.  This  may  be  detected  by  the  odor  of  burning  shellac 
or  rubber.  Dampness  may  produce  the  same  results. 

The  copper  bars  of  the  induction  machine  are  all  short-cir- 
cuited, rendering  attention  to  insulation  unimportant. 

Properties  of  Copper  Wire. — The  manufacturers'  tables  fur- 
nish the  data  for  wires  of  various  diameters  by  which  may  be 
known  their  resistances  to  the  passage  of  a  current.  Below  is  a 
brief  table  which  will  aid  in  the  solution  of  problems  connected 
with  the  wiring  of  a  mine  for  motors. 


Wire  No. 
B.&S.; 
American 
and  Bir- 
mingham 
Gauges. 

Circular  Mils. 

Weight  per 
1000  Feet  of 
Common  Insu- 
lated Wire. 
Pounds. 

Safe  Current- 
carrying  Capa- 
city in  Shaft  or 
Tunnel  Work, 
Amperes. 

Resistances  per 
i  ooo  Feet  in 
Ohms,  at  60°  F. 

B.&S. 

B. 

B.&S. 

B. 

B.&S. 

B. 

B.&S. 

B. 

000 
OO 
0 

I 

a 

4 
6 
8 
10 

12 

167,806 
133.076 
105.534 
83.6Q2 
66,371 
41,741 
26,251 
16,509 
10,383 
6,529 

180,625 
144,400 
115,600 
90,000 
80,686 
56,644 
41,209 
27,225 
17,956 
11,881 

603 
483 
385 
308 
246 
152 

102 
69 

49 
35 

647 

517 
415 

III 
198 
148 

101 

7i 
5i 

295 
250 

210 

180 
J.S0 
no 
80 
60 
40 
30 

315 
272 
227 

193 
161 
118 
85 
64 
43 

.0606 
.0764 
.0964 
.1219 

•1529 
.2446 

•3879 
.6214 

•9785 
1-5520 

.0564 

•°755 
.0881 
.1131 
.1262 
.1797 
.2471 
•3769 
•5670 
•8569 

The  Electric  Symbols  as  applicable  to  motors  are : 
E  =  volts  at  the  terminals; 
V  =  volts  loss  in  transmission ; 
£-|-y=E.M.F.  at  the  generator  terminals; 

/  =  current  required  at  the  motor  to  deliver  H,  mechanical 

horse-power  at  the  shaft  of  the  motor; 
D  =  single  distance  between  motor  and  generator  in  feet; 
L  =  aggregate  length  of  conductor  in  feet; 
fl"=the  number  of  mechanical  H.P.  delivered  at  the  motor 

shaft ; 
A  =the  area  of  cross-section  of  conductor  in  circular  mils; 


ELECTRIC  GENERATION  AND   WATER-POWER.  175 

d=the  diameter  of  the  wire  in  circular  mils; 

R  =  the  total  conductor  resistance  in  ohms  for  the  length,  Ly 
or  zD\ 

W=the  weight  in  pounds  of  copper  conductor; 

m  =  the  commercial  efficiency  of  motor  in  per  cent  ; 

"        "  generator,  90  per  cent; 

S=  "  "        "  whole  circuit,  per  cent; 

p=per  cent  of  energy  lost  in  system. 

Continuous-current  Distribution.  —  The  following  rules  are 
serviceable  as  guides  in  calculating  the  dimensions  of  D.C.  con- 
ductors. The  E.M.F.  varies  directly  with  the  amount  of  energy 
transmitted. 

Knowing  the  work  to  be  done,  line  loss  and  E.M.F.  at  the 
terminals  and  point  of  distribution,  then  the  cross-section  of  the 
conductor  varies  directly  with  the  distance,  and  the  weight,  as 
the  square  of  the  distance. 

The  conditions  remaining  as  given,  the  weight  of  conductor 
will  vary  inversely  as  the  square  of  the  E.M.F.  at  the  terminals. 
Given  the  amount  of  power  which  is  to  be  transmitted  by  a  stated 
weight  of  conductor  and  loss  in  distribution,  then  the  distance 
over  which  the  power  can  be  transmitted  varies  directly  as  the 
E.M.F.  For  a  given  amount  of  power  and  a  given  conductor  it 
varies  as  the  square  of  the  E.M.F. 

The  Resistance  of  Conductors.  —  The  resistance  of  a  mil-foot 
of  pure  copper  at  o°  C.  =  9-59  ohms. 

The  resistance  of  a  mil-foot  of  96  per  cent  conducting  Cu 
(which  is  the  commercial  conductivity  usually  specified  at  70°  F.) 
=  10.81  ohms. 

TT 

Then  the  electrical  horse-power  at  the  motor  terminals  =  —  . 

m 


__    IE          ,     _ 
£T—  —  7    and    /  =  —  5;  —  amperes. 
740  hm 

The  resistance  of  the  conductors,  both  ways,  is 
21.62!)     lo.SoZ, 


176  MANUAL  OF  MINING. 

The  line-drop 

v    IR     2I*2lD    and     A 


Emv       ' 

EXAMPLE.  —  A  motor  of  50  horse-power  of  0.93  efficiency  is  to  be  installed 
in  a  mill  with  the  terminal  volts  at  500.  The  transmission  distance  is  500 
feet,  and  the  allowable  loss  in  that  line  6  per  cent.  Required  the  size  of  con- 

coo 

ductors.    The  E.M.F.  at  dynamo=-  —  =  531.9  volts. 

0.94 

The  line-drop  V=  531.9—  500=31.9  volts,  and  the  area  of  the  conductor 
16,128.5X50X500 


500X0.93X31.9 


27,863  circular  mils. 


The  current  /=  - —  —  =  80.6  amperes. 
500X93 

The  National  Code  allows  only  97  amperes  for  No.  5  B.  &  S.  gauge  con- 
ductor. To  transmit  80.6  amperes  safely,  at  least  No.  4  wire  would  be  re- 
quired. For  underground  work  a  No.  6  wire  would  be  permissib  e. 

The  resistance  of  No.  46.  &  S.  wire  =0.2480  ohms  at  20°  C. 

The  resistance  of  500  feet  of  No.  4  wire  =0.1240  ohms  at  20°  C. 

Then  the  volts-drop  =7^=80.6X0.1240=  09.94  volts;   hence  volts  at  dy- 

OQ.Q 

namo  =  500+ 99.94  =  599-94  volts;  and  per-cent  drop  =• — —  =  1.8 70  per  cent 

599-9 

General   Wiring   Formula   for  Alternating-current  Distribu- 
tion.— 

Let  D= distance  of  transmission  one  way  in  feet; 
W  =  total  watts  delivered  to  consumer; 
p  =  per  cent  of  power,  W,  lost  in  line ; 
E  =  voltage  between  main  conductor  at  the    consumer's 

end  of  circuit; 
K= constant,  whose  value  depends  on  the  kind  of  system 

and  the  power  factor  table,  page  178; 
A  —  factor  for  determining  weight  of  conductor  (table  on 

page  178); 
F= variable,  whose  value  depends  on  the  kind  of  system 

and  the  nature  of  the  load;  and 
M  =  variable,  which  depends  on  the  size  of  wire,  the  power 

factor,  and  the  frequency. 


ELECTRIC  GENERATION   AND   WATER-POWER.  177 


Then  the  area  of  conductor  =  pE2    in  circular  mils. 

Volts  lost  in  line =0.0 1  pEW. 

WF 

Current  in  main  conductor s  =  -g-. 

D*WKA 

Pounds  of  copper  =  —       — ^^. 
i,ooo,oooPE2 

EXAMPLE. — Find  the  size  of  conductor  necessary  to  transmit  40,000  watts  to 
a  distance  of  20,000  feet,  the  line  voltage  being  2000  and  the  permissible  loss 
10  per  cent.  Transmission  is  to  be  by  the  three-phase  three-wire  system, 
at  60  cycles,  power  factor  of  85  per  cent. 

20,000X40,000X1500  .  „ 

Area  of  conductor  =  — ' —  =  30,000  circular  mils. 

10X4,000,000 

The  nearest  size,  No.  5  B.  &  S.  gauge. 

Volts  lost  in  line=o.oiX  10X2000X1.06=  212  volts. 

.T.     40,000X0.68 
Current  in  main  conductors  (/)  =  —  —  =  13.6  amperes. 

Resistance. — In  alternating  current  resistance  is  the  same 
in  kind  as  in  continuous  current,  but  is  generally  considered  as 
negligible  in  comparison  with  other  alternating-current  factors. 
The  effective  value  of  an  alternating  current  is  1.41  times  its 
apparent  value. 

The  Power  Factor  of  an  alternating-current  circuit  is  the 
ratio  of  the  kilowatts,  as  indicated  by  a  watt-meter,  to  the  ap- 
parent watts  or  volt-amperes.  It  enables  one  to  readily  deter- 
mine the  true  energy  in  a  circuit  when  the  apparent  energy  is 
known;  likewise,  the  resistance  when  the  impedance  is  known; 
the  energy  E.M.F.,  when  the  impressed  E.M.F.  is  given,  and 
the  energy  current  when  the  maximum  current  is  known. 

Electric  Systems. — The  available  electric  power  systems  com- 
prise the  continuous-current,  the  polyphase  alternating  current, 
and  a  combination  of  the  two.  When  the  machinery  to  be  installed 
is  merely  traction  and  lighting,  the  continuous-current  system 
will  be  employed.  Being  essentially  a  low-pressure  current,  it 
must  be  used  locally  and  cannot  be  transmitted  over  great  distances. 

Either  a  two-wire  or  a  three-wire  line  may  be  supplied  by 


i78 


MANUAL  OF  MINING. 


Values  of  A,  \ 

as™1- 

Values  of  K, 
•    Per  Cent  Power  Factor. 

Values  of  F, 
Per  Cent  Power  Factor. 

IOO 

95 

QO 

85 

So 

IOO 

95 

90 

85 

80 

Single-phase  
Two-phase  (4-wire).  .  .  . 
Three-phase  (3-wire)..  . 

6.04  2160 
I2.o8|io8o 
9.06  1080 

2400 
1200 
I2OO 

2660 
133° 
1330 

3OOO 
I5OO 
1500 

3380 
1690 
1690 

1.  00 

•5° 
.58 

1-05 

•53 
.61 

i.  ii 

•55 
.64 

1.17 
•59 
.68 

'•*S 
.62 

.72 

VALUES  OF  M  FOR  WIRES  18  INCHES  APART. 


Size  of 
Wire, 

Area  of 
Wire, 

Weight 
of  Wire, 

Resistance 
of  Wire, 
per  1000 

25  Cycles, 
Per  Cent  Power 
Factor. 

40  Cycles, 
Per  Cent  Power 
Factor. 

B  &  S. 

Circular 

Feet  at 

Gauge. 

Mils. 

(in  Lbs.). 

20  C. 

(in  Ohms). 

95 

9° 

85 

80 

9S 

90 

8s 

So 

coo 

211,600 

640.73 

.04879 

•23 

I  .  2C) 

i-33 

•34 

1.58 

•S3 

1.61 

1.67 

00 

133,079 

402.97 

•07758 

•  14 

T.l6 

1.16 

.16 

i.  35 

1.35 

'•37 

4 

41,742 

126.40 

•2473 

.02 

I.  00 

I.  00 

I.  00 

1.05 

.06 

1.03 

I.  00 

5 

33,102 

100.23 

.3120 

.00 

I.  00 

I.  00 

I.  00 

1.03 

I.OI 

I.  00 

I.  00 

6 

26,250 

79-49 

•  3934 

.00 

I.  00 

I.OC 

.00 

I.  01 

.00 

I.  00 

r  .00 

7 

20,816 

63-03 

.4958 

For  sizes  7  and  8  wire  the  constants 

8 

16,509 

49-99 

.6250 

are  within  .05%  (average)  of  5  and  6. 

Size  of 

Area  of 

Weight 
of  Wire 

Resistance 
of  Wire. 

60  Cycles, 
Per  Cent  Power 

1  25  Cycles 
Per  Cent  Power 

Wire, 

Wire, 

per  1000 

Factor. 

Factor. 

B.  &S. 

oare,  per 

Feet  at 

Gauge. 

Mils. 

(in  Lbs.). 

20  C. 

(in  Ohms). 

95 

90 

8s 

So 

95 

90 

8S 

So 

OOO 

211,600 

640.73 

.04879 

1.62 

.84 

i.  00 

2.09 

2-3S 

i.86 

.24 

3-40 

00 

133,079 

402-97 

•07758 

i-34 

•52 

i.  60 

1.66 

1.86 

.18 

.40 

2.57 

5 

33.102 

100.  23 

•2473 
.3120 

i.  08 

.08 

i.  06 

1.04 

I.  21 

•35 
.27 

•3° 

i-43 
1.31 

6 

26,250 

79-49 

•3934 

i  .03 

.02 

I.  00 

I.  00 

I.  12 

.14 

.14  1.13 

I 

20,816 
16,509 

63-03 
49-99 

.4958 
.6250 

For  sizes  7  and  8  wire  the  constants 
are  within  .0^%  (average)  of  5  and  6. 

the  D.C.  generators.  With  the  latter,  the  generator  must  have 
three  wires  to  deliver  continuous  current  of  two  pressures,  one 
being  double  that  of  the  other.  If  this  type  of  dynamo  is  not 
used,  either  two  generators  are  installed  in  series  or  else  some 
complex  form  of  balancing  device  is  required. 

When  the  power  is  required  for  pumping,  lighting,  coal-cut- 
ting, etc.,  the  alternating  current  can  be  more  economically 
transmitted.  Its  machines  are  easier  to  construct  than  D.C. 
machines  of  similar  pressures  and  its  motors  give  less  spark  and 


ELECTRIC   GENERATION  AND   WATER-POWER.  179 

are  of  more  general  application.  Its  higher  efficiency  is  due  to 
the  high  voltage  to  which  it  can  be  carried.  This  is  what  gives 
the  polyphase  current  its  chief  claim.  Particularly  is  this  true 
in  mountainous  districts,  where  sources  of  water-power  distant 
from  the  mines  can  be  developed  for  electric  generation  and 
thence  distributed  to  various  points. 

In  the  event  that  both  systems  must  be  maintained  in  the 
mine,  both  currents  may  be  generated  and  conducted,  or  it  is 
practicable  to  produce  the  cheaper  polyphase  alternating  cur- 
rent, and,  by  the  use  of  converters  at  suitable  points,  lo  change 
only  such  portion  of  the  energy  into  direct  current  as  will  be 
required  for  the  haulage  and  lighting,  leaving  the  remaining 
portion  of  the  current  for  the  other  mining  operations. 

Converters. — Changes  from  the  alternating  current  to  a 
direct  current,  or  the  reverse,  can  be  effected  by  the  means  of  a 
rotary  converter.  Changes  from  a  high-pressure  to  a  low-pres- 
sure current,  or  the  reverse,  whether  D.C.  or  alternating,  may  be 
effected  by  a  transformer  for  stepping  down  or  stepping  up  in 
pressure.  As  reliability  is  of  the  utmost  importance,  transformers 
are  always  installed  of  twice  the  capacity  for  the  expected  service. 

Their  efficiency  is  about  96  per  cent. 

The  Transmission  Difficulties. — The  great  objection  to  elec- 
tricity is  the  high  cost  of  wire  and  erection.  Economy  of  copper 
wire  is  possible  only  by  employing  high  voltage  in  the  currents, 
as  shown  by  the  formulae  already  discussed.  As  an  example, 
195  Ibs.  of  copper  can  transmit  50  amperes  at  2000  volts  over  a 
5ooo-foot  line  with  the  same  percentage  of  line  loss  that  795  Ibs. 
of  No.  36.  &  S.  gauge  can  transmit  the  same  quantity  at  500 
volts.  But  in  the  transmission  other  difficulties  are  introduced. 
The  insulation  must  be  perfect  and  the  wires  separated  from  one 
another  sufficiently  to  prevent  an  induction  of  one  current  upon 
the  other.  From  both  sources  the  line  loss  would  be  large. 
With  proper  insulation,  however,  the  limit  of  voltage  which 
can  be  carried  would  be  fixed  by  the  distance  between  the 
wires.  There  must  be  a  wide  gap  between  them,  for,  with  the 
enormous  voltage  which  is  now  being  used,  we  are  approaching 


l8o  MANUAL  OF  MINING. 

the  conditions  in  lightning,  not  in  a  single  stroke  but  a  number 
of  strokes  following  each  other  in  very  rapid  succession.  The 
length  across  the  air-space  increases  as  the  voltage  increases,  and 
for  currents  of  50,000  volts  the  arms  carrying  the  wires  at  their 
ends  must  be  12  feet  long. 

There  is  no  limit  to  the  E.M.F.  which  the  alternating-cur- 
rent machines  can  impart  to  the  current.  The  majority  of  long- 
distance transmission  companies  employ  the  three-phase  system 
and  at  as  high  a  difference  of  potential  as  possible.  A  high- 
voltage  current  may  be  obtained  either  directly  at  the  dynamo 
whence  it  is  delivered  to  the  distant  end,  or  the  dynamos  may 
generate  a  low-tension  current  which  is  transformed  into  the  high- 
voltage  current  at  the  dynamo  station  for  transmission  to  the 
distant  point.  At  the  mine  it  may  be  transformed  wholly  into 
alternating  current  of  tension  suitable  for  the  machines,  or  only 
in  part,  the  balance  being  converted  to  a  continuous  current  for 
haulage. 

The  Most  Economical  Area  of  Wire. — The  losses  in  trans- 
mission are  due  either  to  resistances  or  to  leakage.  The  drop 
in  pressure  caused  by  leakage  of  the  current  from  the  wire  through 
the  points  of  support  is  deduced  by  insulating  the  supports  and 
an  enveloping  tube  of  non-conducting  material  around  a  rubber- 
covered  wire.  The  drop  in  voltage,  due  to  interaction  of 
neighboring  parallel  circuits,  is  averted  by  an  ample  air-gap. 
That  due  to  the  resistance  of  the  wire  to  the  passage  of  the  cur- 
rent is  controlled  by  the  use  of  a  larger  wire.  The  resistance  of 
the  wire  varies  inversely  as  the  area  of  its  cross-section.  A  small 
wire  offers  a  great  resistance  to  the  electric  fluid.  This  loss 
reduces  the  available  pressure  at  the  terminals,  and  consequently 
the  amount  of  power  there  obtained.  It  develops  heat  in  the  con- 
ductor, which  is  not  only  injurious  to  the  wire,  but,  without 
secure  fire  protection,  is  also  dangerous  to  the  surrounding  medium. 

In  calculating  the  dimensions  of  the  conducting  wire  a  balance 
is  made  between  the  elements — great  first  cost  of  a  large  quantity 
of  copper  and  the  waste  of  power  during  the  life  of  the  plant  by 
use  of  small  wire.  The  allowable  expenditure  of  wire,  in  any 


ELECTRIC  GENERATION  AND   WATER-POWER.  181 

given  plant,  will  determine  the  E.M.F.  which  can  be  used;  the 
efficiency  required  of  the  plant  will  fix  the,  allowable  line  loss. 
Hence  the  most  economical  area  of  conductor  is  that  for  which 
the  annual  interest  on  the  capital  outlay  equals  the  annual  cost  of 
the  energy  thus  wasted. 

The  Allowable  Voltage  in  Mines. — There  are  many  con- 
flicting requirements  for  electric  wiring  in  mines.  The  great 
distances  of  distribution  require  high  voltage  to  save  copper. 
This  demands  a  high  degree  of  insulation,  which  is  difficult  to 
maintain  in  the  moist  air  and  the  mine-gases.  Moreover,  the 
conditions  of  work  require  bare  wires,  and  their  position  is  such 
as  make  them  dangerous  to  life.  The  porcelain  or  glass  tubes 
are  convenient  insulators,  but  they  do  not  prevent  shock. 

The  lowest  difference  of  potential  causing  fatal  accidents  is 
220  volts  with  a  direct  current  and  no  volts  with  an  alternating 
current.  The  more  the  difference  of  potential  increases  beyond 
these  figures  the  more  must  all  tensions  be  regarded  as  equally 
dangerous.  Though  a  higher  voltage  would  be  more  economical 
in  the  cost  of  copper  for  a  given  power  transmission,  the  miner  is 
restricted  to  a  low  voltage  because  of  the  risks  imposed  by  fire  and 
shock.  The  allowable  maximum  voltage  "has  been  fixed  at  450. 

In  the  three-wire  system  one  trolley  wire  is  used  on  each  side 
of  the  traction  generator  or  to  the  free  sides  of  two  generators 
in  series.  In  the  former  case  the  conductors  from  the  rails  should 
connect  with  a  small  dynamo  to  keep  the  rail  pressure  to  one-half 
that  of  the  main  terminal.  In  the  latter  case  all  rail  conductors 
are  taken  to  a  common  point  of  connection.  By  grounding  the 
neutral  wire  the  shock  at  contact  would  not  be  so  serious  unless 
both  wires  were  simultaneously  touched.  The  trolley  circuits, 
however,  subject  the  maker  of  the  contact  to  the  full  discharge 
of  pressure  between  the  two  sides  of  the  three-wire  system.  So, 
too,  a  ground  connection  of  the  high-pressure  circuits,  or  a  cross 
between  the  high-  and  low-pressure  circuits,  inside  the  trans- 
former, would  subject  him  to  fatal  results. 

For  traction  a  pressure  of  250  volts  is  qtiite  sufficient,  even 
with  an  induction  motor,  for  i;  can  be  operated  by  transformers 


1 82  MANUAL  OF  MINING. 

at  every  few  hundred  feet  to  keep  the  voltage  down.  Incan- 
descent lamps  on  a  two- wire  D.C.  system  are  operated  on  125 
or  250  volts,  and  at  double  these  pressures  on  the  three-wire 
distribution  system.  On  alternating-current  circuits  the  lamps 
are  supplied  by  near-by  transformers.  Lamps  and  stationary 
motors  are  carried  under  a  maximum  pressure  of  250  or  500 
volts,  with  corresponding  differences  of  pressure  between  any 
wire  and  the  earth  of  one-hal?  the  quantity. 

The  Motor. — A  motor  converts  the  current  into  mechanical 
energy.  It  is  directly  connected,  or  is  belted,  to  the  mechanism 
to  be  operated.  The  construction  of  the  motor  is  practically  the 
same  as  the  dynamo,  but  its  function  is  exactly  the  opposite.  The 
current  flows  in  the  reverse  order,  passing  through  the  coils  of 
the  armature  and  the  field-magnets.  These  react  upon  the  coils, 
carrying  a  current  into  the  armature  and  thus  produce  rotation. 
The  force  turning  the  armature  is  called  its  torque  and  is  its 
capacity  to  do  work.  Around  the  field,  magnetized  by  the  cur- 
rent flowing  around  the  magnets,  a  current,  called  the  back  E.M.F., 
is  induced  in  the  opposite  direction.  This  increases  with  the 
current  intensity  and  the  number  of  coils  and  their  speed.  Taking 
current  in  proportion  to  the  work  performed,  the  motor  is  self- 
acting. 

The  continuous  motors  are  wound  in  three  ways,  correspond- 
ing to  those  of  the  dynamo,  each  having  its  own  field  of  useful- 
ness. In  the  series  motor  there  is  but  one  continuous  circuit 
through  the  stationary  and  rotary  pieces.  This  exerts  a  greater 
starting  torque,  but  its  speed  will  change  with  every  variation 
of  the  load.  It  is  intended  for  constant  load  only.  If  it  is  over- 
loaded, its  speed  is  slackened.  Its  torque  is  increased  by  its 
greater  draft  of  current,  and  if  it  is  overloaded,  it  runs  slower, 
the  field-magnets  heat  up  and  burn  out,  or  a  fuse  blows.  Re- 
moving the  load  suddenly  causes  it  to  race,  because  the  field 
is  weakened  and  the  back  E.M.F.  is  diminished.  The  increased 
speed  which  ensues  may  ultimately  injure  the  motor.  The  shunt 
motor  will  maintain  a  constant  speed  and  a  back  E.M.F.  even 
with  variations  of  the  load,  but  its  starting  power  is  low.  Designed 


ELECTRIC  GENERATION  AND   WATER-POWER.  183 

for  the  maximum  load  which  it  is  expected  to  drive,  it  is  employed 
in  mining  work.  It  is  heavier  than  the  series  motor.  The 
compound  motor  is  used  about  the  mines  and  combines  the 
shunt  and  the  series  features  of  motors. 

Alternating-current  Motors. — These  motors  are  either  syn- 
chronous or  non- synchronous.  The  speed  of  the  former  type 
bears  a  constant  relationship  to  the  frequency  of  the  driving 
current.  Being  unable  to  start  until  it  synchronizes  with  that 
of  the  generator,  miners  use  it  very  little. 

In  the  non-synchronous  motor,  also  called  an  induction"  motor, 
the  speed  varies  with  the  load  and  is  independent  of  the  frequency 
of  the  current.  It  is  self -starting.  Induction  motors  may  be 
obtained  of  the  two-phase  four-wire  type,  or  the  polyphase  type 
with  three  wires.  In  the  two-phase  motors  each  coil  of  a  set  in  the 
rotor  receives  its  own  current.  One  induces  a  current  in  the 
conductors  of  the  rotor,  causing  attraction  between  them  and 
the  stator  coils.  The  slight  rotation  that  follows  changes  the 
current  to  the  next  set.  Each  set  of  windings  produces  similar 
results  and  adds  to  the  rotary  effort.  In  the  three-phase  system 
continuous  rotation  occurs  under  similar  conditions,  except  that 
the  entering  current  is  divided  among  three  separate  sets  of  coils. 
A  back  E.M.F.,  as  an  opposing  pressure,  is  induced  in  the  con- 
ductors of  the  rotor,  as  in  the  D.C.  machine.  When  the  load 
is  increased  the  speed  falls,  the  opposing  pressure  becomes  less. 
There  is  an  increase  in  the  available  driving  pressure,  or  torque, 
and  an  increase  of  current  in  the  rotor.  This  alters  its  speed 
relative  to  the  rotating  field.  Every  alteration  of  current  in  the 
rotor  conductors  reacts  on  the  stator  coils  and  reduces  the  strength 
of  the  magnetic  field,  until  the  energy  supplied  to  the  motor  is 
insufficient  for  the  work  and  the  motor  stops.  In  this  respect 
the  induction  motor  differs  from  the  D.C.  motor.  An  increase 
of  load  brings  the  former  to  a  standstill,  but  results  in  the  burning 
of  armature  coils  in  the  latter.  An  induction  motor  can  take 
an  excess  of  20  per  cent  over  its  normal  load  without  injury. 
Having  neither  commutator,  induction  brushes,  nor  movable 
contacts,  it  is  the  ideal  motor  for  mining  machinery  where  the 


1 84  MANUAL  OF  MINING. 

air  may  be  laden  with  dust,  gas,  or  moisture.    The  power  factor 
is  over  90,  with  an  efficiency  proportionately  high. 

The  objections  urged  against  the  polyphase  induction  motoi 
are  its  excessive  starting  current,  which  results  in  a  drop  of  voltage 
with  consequent  bad  regulation  to  other  apparatus  in  current, 
and  its  inflexibility  in  the  matter  of  speed  regulation.  The  first 
objection  against  the  short-circuited  secondary  type  is  negligible 
if  the  power  be  increased  at  the  generating  station  on  starting. 
The  machine  is  essentially  a  constant- speed  motor.  When 
variable  speed  is  required  it  would  be  more  advantageous  to 
employ  a  variable  resistance  with  the  secondary  type  of  induction 
motor  effected  by  a  rheostat. 

Rheostats. — There  is  no  back  E.M.F.  in  the  circuits  at  the 
moment  of  starting  a  motor,  and,  to  prevent  the  heavy  entering 
currents  from  doing  damage  to  the  armature,  extra  resistances 
must  be  introduced  into  the  current  until  such  times  as  the  motor 
has  got  up  speed.  The  starting  resistance  consists  of  coils  of 
German-silver  or  iron  wire  contained  in  fire-proof  boxes.  A  mova- 
ble contact,  turned  to  one  side  or  the  other  from  its  central  position, 
determines  the  direction  of  rotation.  It  is  operated  by  an  attend- 
ant.' On  closing  the  circuit  the  current  flows  through  the  entire 
resistance  of  all  the  coils.  As  the  motor  speeds  up,  one  after 
the  other  of  the  coils  in  the  circuit  is  gradually  cut  out.  The 
diverter  of  columns  of  iron  and  mica  in  five  separate  portions  is 
used  for  locomotives. 

The  resistance,  on  shunt-  and  compound-wound  machines,  is 
placed  in  the  shunt  winding  of  the  fields  to  increase  or  decrease 
the  E.M.F.  There  is  usually  on  pump-motors  an  additional 
safeguard  of  an  automatic  release  for  the  rheostat  contact  lever. 
The  contact  arc  is  worked  against  a  spring,  and  is  held,  when 
fully  on,  by  means  of  a  magnet,  which  is  connected  in  series  with 
the  shunt  fields  of  the  motors. 

For  alternating-current  machines  the  starting  resistance  con- 
sists of  two  rings  connected  to  the  conductors  and  insulated  from 
one  another  and  from  the  motor  shaft.  Fitted  to  them  are  brushes 
with  a  resistance  inserted  between  them,  which  can  be  automatic- 


ELECTRIC   GENERATION   AND   WATER-POWER.  185 

ally  cut  out  as  the  motor  attains  its  speed.  These  slip- rings 
are  covered  in  mining  machines  to  confine  the  spark.  Another 
class  of  starter  is  that  in  which  the  current  passes  through  water 
into  which  is  lowered  a  lead  cone  from  the  surface  to  the  bottom 
where  the  other  terminal  is  located.  This  form  of  switch  is  free 
from  spark,  though  it  might  be  dropped  too  quickly  and  thus  be 
almost  valueless. 

The  variable  speed  induction  motor  has  a  two- drum  con- 
troller for  resistance  and  reversing  operated  by  a  single  handle. 
The  resistance  is  in  series  with  the  rotor  windings,  and  is  cut  out 
in  the  usual  way  until  the  approximately  constant  speed  is  attained. 

Reversing-motors. — When  the  direction  of  the  motor  is  to  be 
reversed,  its  brushes  should  bear  vertically  on  the  commutator 
and  a  re  versing- switch  be  used.  This  appliance  for  changing  the 
direction  of  the  current  in  the  field-magnets  is  arranged  as  a  part 
of  the  starting  resistance. 

Enclosed  Motors. — If  possible,  motors  should  be  run  open, 
for  if  totally  enclosed  a  much  larger  motor  will  be  required  for  the 
same  work  on  account  of  the  difficuLy  of  dissipating  the  heat 
developed  in  the  coils.  If  precautions  are.  observed  not  to  over- 
load the  machine,  the  danger  is  reduced  and  certainly  would  not 
be  greater  than  that  existing  from  either  cause  in  the  daily  opera- 
tions. They  are  rarely  placed  near  gaseous  mixtures,  and  the 
danger  of  igniting  neighboring  timbers  by  their  spark  is  very 
slight.  For  safety  in  mines,  however,  electric  motors  are  enclosed 
to  protect  them  from  fall  of  rock,  dust,  and  dampness. 

The  Efficiency  of  a  Motor. — Not  all  the  current  which  the 
motor  receives  is  usefully  applied.  Part  of  it  is  spent  in  heating 
the  circuit  of  the  motor  and  part  in  overcoming  frictional  resist- 
ance. The  remainder,  which  is  available  for  mechanical  work, 
is  still,  however,  quite  large.  As  large  motors  have  high  efficiency, 
it  is  possible  to  attain  very  high  results,  if  their  capacity  be  large 
enough.  This  is,  however,  a  question  of  weight  and  first  cost. 
The  efficiency  of  a  machine  is  not  far  from  90  per  cent  under 
average  conditions,  and  will  reach  92  per  cent  when  operating 
under  the  load  and  speed  for  which  it  was  designed.  The 


i86 


MANUAL  OF  MINING. 


table  below  gives  the  number  of  watts  necessary  to  produce  one 
effective  horse-power  at  the  terminals  of  a  motor  with  stated 
commercial  efficiencies.  Thus  a  motor  with  10  effective  horse- 
power having  an  efficiency  of  85  per  cent  will  require  8780  watts 
of  current. 


Efficiency 
of  a  Motor. 

Watts  to  be 
Delivered  to  it 
per  Horse-power. 

T.OO 

0.98 

746 
764 

o-9S 

774 

0.90 
0.85 
0.80 

830 
878 
933 

o-75 

1000 

EXAMPLE. — The  efficiency  of  a  50-H.P.  motor  is  0.90.  It  is  to  be  driven  a 
half  mile  from  a  dynamo  whose  efficiency  is  0.85.  The  modulus  of  the  engine 
driving  the  generator  is  0.80.  Required  th  horse-power  of  the  engine. 
Volt  ge,  400. 

The  motor  requires  830  watts  per  horse-power  or  41,500  watts  total: 

41,500-^400=103.75  amperes. 

For   this   the  wire  must  be  No.  4.     Its  resistance,  per  thousand  feet,  is 
0.245  ohms. 

The  watts  lost  per  thousand  feet  are 

PR=  104  X  104X0.245=  2650. 

For  the  mile  of  line  in  the  circuits  13,992  watts  are  lost. 
The  dynamo  must  therefore  furnish 

41,500+13,992=54,492  watts. 

This  corresponds  to  a  voltage  at  the  dynamo  of  533,  and  a  <oss  of  33  per  cent 
55,492 


B.H.P= 


=  87.6      nd 


746X0.85 

A  No.  2  wire,  B.  &  S.  wire,  will  show  a  loss  of  20  per  cent,  a  dynamo  voltage 
of  481,  and  an  I.H.P.  of  100. 

Water-power. — In  the  immediate  vicinity  of  the  wheel  water 
the  power  has  been  employed  for  various  operations,  but  not 
until  the  modern  successful  utilization  of  electricity  has  a 
cheap  efficient  means  been  discovered  for  its  transmission  to 
great  distances.  The  installation  of  electric  plants  is  opening 
the  possibilities  of  water-power  to  an  enormous  degree. 


ELECTRIC   GENERATION  AND   WATER-POWER.  187 

The  gross  power  possessed  by  water  and  the  potential  energy 
which  is  capable  of  being  transmitted  and  converted  into  kinetic 
energy  is  measured  by  the  product  of  the  weight  of  water  dis- 
charged by  the  height,  h,  through  which  it  has  fallen.  Assuming 
the  weight  of  a  cubic  foot  of  water  at  62.5  and  Q  as  the  number 
of  cubic  feet  discharged  per  minute,  then  a  horse-power  equals 
o.ooi6i<2&.  The  net  power  obtained  from  this  in  effective  work 
is  from  40  to  90  per  cent  according  to  the  character  of  wheel 
used. 

This  energy  develops  power  by  driving  one  of  several  types 
of  wheels,  of  which  the  undershot  and  the  overshot  were  the 
earliest.  They  consume,  however,  enormous  volumes  of  water 
and  are  used  only  for  small  machinery.  The  turbine-wheels, 
placed  horizontally,  revolving  under  a  pressure  due  to  a  head 
of  water  above,  give  a  high  efficiency  with  the  use  of  a  large 
volume  of  water.  These  wheels  are  not  very  large  and  are  en- 
cased in  a  globe  or  cylindrical  casing  above  which  is  a  penstock 
through  which  the  water  flows.  The  water  enters  centrally 
and  discharges  circumferentially  and  produces  a  high  rate  of 
revolution.  The  heads  at  which  these  turbines  are  used  rarely 
exceed  200  feet. 

Impulse  Wheels. — The  more  recent  forms  of  wheels  are  of 
the  impulse  type,  in  which  a  small  wheel  receives  at  its  circum- 
ference the  impact  of  a  stream  flowing  with  a  high  velocity  due 
to  a  head  which  in  some  cases  is  as  great  as  1900  feet.  The 
periphery  of  the  wheel  carries  a  number  of  small  cup-shaped 
vanes  whose  curves  and  position  are  such  as  to  receive  the  full 
impact  of  the  stream  and  to  discharge  the  water  at  a  velocity  of 
nearly  zero.  The  efficiency  of  these  wheels  is  therefore  high, 
and  their  construction  is  very  simple ;  the  inferior  limit  of  head 
with  which  they  may  operate  to  advantage  is  about  30  feet.  The 
maximum  limit  of  head  is  determined  only  by  the  strength  of 
the  material  in  the  wheel.  As  is  the  case  with  the  De  Laval 
steam-turbine  (Fig.  48),  multiple  jets  are  also  used  on  water-tur- 
bines. 

The  diameter  of  the  wheel  is  proportioned  to  the  rate  of 


1 88  MANUAL  OF  MINING. 

revolution  desired  for  the  main  shaft.  Usually  the  latter  is  that 
for  which  the  electric  generator  has  been  wound  and  designed. 
The  diameters  vary  between  18  and  90  inches. 

There  are  several  types  of  impulse  motors  in  America,  of 
which  the  Doble,  Pelton,  and  Knight  are  excellent  patterns. 
The  efficiency  of  these  wheels  when  properly  regulated  by  gov- 
ernors is  85  per  cent  of  the  theoretical  head,  due  to  the  velocity 
of  discharge  from  nozzles  against  the  wheel-cups.  In  the  moun- 
tainous districts,  where  the  numerous  creeks  do  not  carry  much 
water  but  have  a  large  fall,  these  wheels  may  be  used  to  advan- 
tage by  laying  a  pipe-line  from  the  wheel  to  some  elevated  reser- 
voir in  the  creek  constructed  for  the  purpose.  Not  infrequently 
two  or  even  three  nozzles  feed  the  wheel  when  the  quantity  is 
large  enough  to  permit  it. 

The  supply  of  water  delivered  to  the  wheel  is  usually  esti- 
mated in  miners'  inches,  representing  a  flow  of  something  near 
1.5  cubic  feet  per  minute.  The  miner's  inch  is  known  as  the 
volume  of  water  which  can  be  discharged  through  each  square 
inch  of  an  aperture  2  inches  high  and  4  inches  long  which  is  cut 
through  a  plank  1.25  inches  thick,  the  lower  edge  of  the  aper- 
ture being  2  inches  above  the  bottom  of  the  measuring-box  and 
the  upper  edge  5  inches  below  the  level  of  the  water. 

Flow  of  Water  through  Pipes. — Owing  to  the  roughness  of 
the  internal  surface  of  a  pipe  and  the  restrictions  which  occur  at 
elbows,  joints,  and  valves,  *the  actual  velocity  of  discharge  of 
water  from  a  pipe  is  not  equal  to  that  due  to  the  pressure  at  its 
inlet.  Any  change  of  area  or  of  direction  affects  the  flow  of  the 
water  and  produces  frictional  resistances  which  are  directly  pro- 
portional to  the  length,  inversely  as  the  diameter,  and  increase 
with  the  velocity.  Though  a  theoretical  formula  may  be  evolved 
which  expresses  the  relation  between  the  discharge  of  water  and 
the  head  producing  such  discharge,  empirical  formulae  are  em- 
ployed with  coefficient  inserted  to  provide  for  the  various  elements. 

The  formulae  below  are  given  to  assist  the  engineer  in  deter- 
mining the  dimension  of  a  pipe  requisite  for  the  given  flow  of 
water  with  a  stated  resistance. 


ELECTRIC   GENERATION   AND   WATER-POWER, 


189 


Fluid  Friction  in  Pipes. — Assuming  very  long  clean  pipes  of 
uniform  size,  the  resistance  which  is  afforded  by  the  interior 
surface  of  the  pipe  to  flow  is  measured  in  terms  of  the  head. 


Let  Q  =the  quantity  of  water  discharged,  cubic  feet  per  second; 
v=the  velocity  in  feet  per  second; 


190  MANUAL  OF  MINING. 

d=ihe  diameter  of  the  pipe  in  feet; 
/=the  length  of  pipe  in  feet; 
/=the  coefficient  of  friction; 
Jf=the  gross  head  of  water — the  difference  in  elevation 

between  the  two  termini  of  the  pipe; 
h  =the  frictional  loss  measured  in  terms  of  feet  of  head,  and 
ht  =the  effective  head  =H—  h. 

In  clean  pipes  of  smooth  bore,  /  =0.004  and  in  ordinary  mine 
pipes  /  =0.006. 


>.55S 
| 


Qi.802 

h  =0.000606/^857 


/)i.8o2-io.aoo 

d  =  \  o.ooo6o6/ —  • 

L  h   J 

Q,  for  the  maximum  horse-power, 


o(.ooi70°-5SS 
The  maximum  horse-power  to  be  obtained  is  equal  to 

H.P.=: 


In  Fig.  73  are  curves  based  upon  the  above  formulae,  giving 
the  loss  of  head  due  to  the  frictional  flow  through  pipes  which 
are  ordinarily  used  in  pumping  and  for  transmitting  power  to 
impulse  wheels.  On  the  right-hand  side  will  be  found  a  verti- 
cal line  giving  the  theoretical  horse-power  necessary  to  raise  a 
given  quantity  of  water  indicated  on  the  left  vertical  line  through 
a  height  of  100  feet.  Thus,  raising  7  cu.  ft.  per  second  through 
any  pipe  for  a  height  of  100  feet  requires  80  horse-power  theo- 
retically. When  the  height  is  other  than  100  feet  the  result 
obtained  from  the  diagram  is  a  multiple  of  the  loo-ft.  eleva- 


ELECTRIC   GENERATION   AND   WATER-POWER. 


191 


tion.  To  raise  the  same  quantity  of  water  in  the  same  pipe 
through  600  feet  of  pipe  will  therefore  require  (80X6  =  )  480 
horse- power 

FEET  OF  HYDRAULIC  HEAD  LOST  PER  1,000  LINEAR  FEET  OF  PIPE. 


FRICTIONAL  HORSE-POWER  PER  1,000  LINEAR  FEET  OF  PIPE. 

FIG.  73. — Loss  of  H\draulic  Head  in  Pipes. 

The  horizontal  line  at  the  top  gives  the  amount  of  head  lost 
by  friction,  and  the  horizontal  line  at  the  bottom  the  horse-power 
needed  to  overcome  the  friction.  Thus  let  it  be  required  to  find 
the  head  lost  and  the  horse-power  while  raising  7  cu.  ft.  pei 


IQ2  MANUAL   OF  MINING. 

second  in  a  1 2-inch  pipe  5000  feet  long  on  a  slope  of  12  per  cent. 
The  curve  shows  that  1000  feet  of  1 2-inch  pipe  carrying  this 
volume  of  water  consumes  20  feet  of  head.  Thus  5000  feet  will 
consume  100  feet  of  the  head.  Likewise  1000  feet  will  consume 
in  this  lost  head  about  28  horse-power,  while  5000  feet  will  con- 
sume 140  horse-power  in  overcoming  friction. 

The  frictional  loss  in  a  lo-inch  pipe  1000  feet  long,  carrying 
2,ico,ooo  gallons  per  day  (3.245  cu.  ft.  per  second),  is  12.8  feet. 
An  8-inch  pipe  would  have  consumed  38.5  feet  of  head  under 
the  same  conditions  of  length  and  discharge. 

EXAMPLE. — A  pipe  500  feet  long  is  to  deliver  4  cubic  feet  per  second.  Re- 
quired its  diameter  if  its  frictional  loss  of  head  allowed  is  30  feet.  For  30  feet 
hydraulic  head  per  thousand  feet  of  pipe  and  4  cubic  feet  of  flow  the  diameter 
should  be  9  inches.  500  feet  of  pipe  consuming  30  feet  will  correspond  to 
a  length  of  1000  feet  consuming  60  feet  of  head.  For  this  condition  and 
4  cubic  feet  of  discharge  the  pipe  may  be  of  8  inches  diameter. 

Kutter's  Formula  for  Ditches  and  Flumes. — Fig.  74  contains 
curves  plotted  according  to  the  Kutter  formula,  which  is  accept- 
able for  all  ditches  and  flumes  of  medium  dimensions.  The 
formula  is  based  on  the  average  condition  of  the  frictional  sur- 
face. The  upper  curves  are  for  flumes  with  hydraulic  radii  of 
from  i  to  9.  The  coefficient  of  roughness,  N,  for  flumes  is  taken 
at  0.01 1,  and  for  ditches  at  0.026.  On  the  left  are  laid  off  the 
gradients  in  terms  of  length  of  line  giving  a  fall  of  i  foot,  and  the 
computed  velocity  of  flow  is  on  the  top  line.  Then  the  four 
lower  curves  on  the  left  represent  the  relation  between  velocity 
and  grade  in  ditches  having  hydraulic  mean  radii,  H.M.R.,  of  0.5, 
i,  2,  and  3,  respectively.  By  the  hydraulic  mean  radius  is  under- 
stood the  quotient  obtained  by  dividing  the  cross-sectional  area 
of  the  ditch  by  the  length  of  the  watered  surface  of  the  ditch. 

EXAMPLES. — i.  What  would  be  the  loss  of  head  in  pumping  2000  gallons  of 
water  per  minute  through  an  8-inch  pipe  600  feet  high?  (^  =  4.44,  and 
^=47.6,  or  38.3  feet,  according  to  the  equations  employed. 

2.  What  horse-power  is  consumed  in  overcoming  friction  in  the  previous 
example?  Assume  h  to  be  38.3.  19.4  H.P. 

The  flow  is  4.44  cubic  feet  per  second,  and  the  horse-power  is  0.1134  QH. 


ELECTRIC   GENERATION   AND   WATER-POWER. 


193 


3.  What  horse-power  will  be  given  out  by  the  discharge  of  400  cubic 
feet  of  water  per  minute  from  a  pipe  of  13  inches  diameter,  2600  feet  long, 
•with  a  head  of  400  feet? 

The  loss  of  head  is  32.53  feet,  and  the  horse-power  available  is  276. 

VELOCITY,  FEET  PER  SECOND. 


\  ELOCITY,  FEET  PER  SECOND. 


FIG.  74. — Flow  of  Water  in  Flumes  and  Ditches,  according  lo  Kutter's  Formula. 

4.  Let  it  be  required  to  determine  the  flow  in  a  trapezoidal  ditch  whose 
area  is  72  square  feet  and  whose  wetted  perimeter  is  24.  The  H.M.R.  is  there- 
fore three.  If  the  fall  of  the  ditch  is  i  foot  in  1500  feet,  the  velocity  of  the 


194  MANUAL  OF  MINING. 

flow  is  3.5  feet  and  the  discharge  is  (72X3.5=)  252  cubic  feet  per  second. 
What  should  be  the  grade  of  a  flume  ia'X6'  which  has  a  H.M.R.  of  3,  when 
the  quantity  desired  is  to  be  547  cubic  feet  per  second?  This  corresponds 
to  a  velocity  of  discharge  7.58  feet  per  second. 

In  the  event  that  a  ditch  under  computation  has  a  mean  hydraulic  radius 
other  than  that  platted  among  the  curves  it  may  be  easily  interpolated.  Thus 
one  of  an  h.m.r.  of  ij  on  a  grade  of  i  foot  in  400  feet  will  be  found  to  have 
a  velocity  of  about  3.8  feet  per  second. 

5.  A  pipe  is  500  feet  long  and  3  inches  diameter.  What  should  be  the 
head  to  produce  a  discharge  of  180  feet  per  minute? 

Here  Q=3,  1  =  500,  ^=0.25,  and  assuming  /  to  be  0.00566, 

h=  0.1007  X  0.00566  X  500  X  9  X  1024=  2624  feet. 


H=h  +  7^  =  2624  +  —  =3124  feet. 

6.  What  diameter  should  it  have  to  deliver  the  same  quantity  of  water 
with  a  head  of  82  feet? 

7.  Required  the  flow  of  water  through  a  pipe  2000  feet  long,  13  inches 
in  diameter,  and  200  feet  head.     For  the  maximum  horse-power  we  have 

(20o)°-tM(i.o8)*-tM      T       =0.864  cubic  feet  per  second. 

REFERENCES. 

Electric  Power  in  Mines  (Economy),  Carl  Pfanauch,  Am.  Mfr.,  May  i, 
1896;  Electric-power  Appliances  in  the  Mines  of  Europe,  Emile  Guarini, 
Eng.  Mag.,  Sept.  1003  and  Aug.  1903. 

Enumeration  of  Electric  Plants  in  Rocky  Mountain  Region,  I.  Hale,  Amer. 
Inst.  M.  E.,  1897;  Electric  Mining  in  the  Rocky  Mountain  Region,  J.  Hale, 
A.  I.  M.  E.,  Vol.  XXVI,  1071  and  402;  Enumeration  of  Electric  Plants  in 
Colorado,  Colorado  S.  &  M.  Quart.,  Vol.  I,  1892,  56;  in  Rocky  Mountain 
Region,  by  I.  Hale,  A.  I.  M.  E.,  1897;  Electric  Mining  Machinery  in  the 
British  Collieries,  Sidney  F.  Walker,  Eng.  Mag.,  Aug.  1900. 

Electrical  Development,  Practical  and  Impossible,  William  Baxter,  Eng. 
Mag.,  Oct.  1806,  113. 

Electric  Power  Transmission  in  Mining,  A.  Bloemendal,  Stahl  &  Eisen, 
Nov.  15,  1899;  and  Long-distance  Transmission,  R.  Hutchinson. 

Electric  Transmission  in  Mining,  H.  C.  Spaulding,  Amer.  Inst.  M.  E., 
XIX,  258;  M.  B.  Holt,  Amer.  Inst.  M.  E.,  XX,  316;  F.  O.  Blackwell,  Amer. 
Inst.  M.  E.,  XXIII,  400;  Electric  Transmission  Lecture,  Colliery  Manager, 
1894,  35;  Siemens,  Trans.  M.  &  M.  Eng.,  XLIV,  Part  II,  205;  Tests  of 
Condensing  Engine  and  Generators,  Trans.  M.  &  M.  Inst.,  XLIV,  Part  IL,. 
207;  Electric  Problems  in  Mining,  Coll.  Eng.,  Dec.  1856,  217. 


ELECTRIC   GENERATION   AND   WATER-POWER.  195 

The  Dangers  of  Electricity  in  Mining,  Gluckauf,  Jan.  30,  1904. 

Electric  Motors,  Lecture,  Colliery  Manager,  1894,  34;  Electricity  VS. 
Compressed  Air  in  Colorado,  L.  Searing,  Elec.  Eng.,  Nov.  1896,  528. 

The  Construction  of  Small  Water  Motors,  Elec.  Lond.,  Mar.  9,  1904; 
Water  Wheels,  General,  Knight  Catalogue,  Tables  of  H.P.;  Min.  Bureau 
of  California,  i3th  Report:  Regulators,  Water  Wheels. 

The  Testing  Water  Wheels,  Elect.  Power  &  Gas  Jour.,  Vol.  XII,  32; 
The  Tangential  Water  Wheel,  W.  A.  Noble,  A.  I.  M.  E.,  Vol.  XXIX,  852. 

Electric  Power  in  European  Collieries,  Cass.  Mag.,  July  1904,  243;  Long- 
distance Transmission,  Cass.  Mag.,  July  1904,  112;  Distribution  of  Elec- 
tricity, Cass.  Mag.,  July  1904,  180. 

Flow  of  Water  in  Pipes,  Proc.  Am.  Soc.  C.  E.,  XXIX,  821. 


CHAPTER  VH. 

HOISTING  MACHINERY  AND  UNDERGROUND  CONVEYANCES. 

Underground  Conveyances. — Where  possible,  the  mineral  is 
loaded  at  the  face  into  cars  or  buckets  to  avoid  frequent  handling. 
But  in  metal-mines  and  in  steep-pitching  coal- veins  this  is  not 
possible,  and  the  product  is  delivered  through  shutes,  or  bat- 
teries, to  the  cars  at  the  lower  level. 

Where  the  output  is  small  and  the  depth  of  shaft  is  slight, 
metal-mines  employ  buckets  which  are  frequently  drawn  through 
the  haulage-way  on  the  trucks,  and  thence  hoisted  to  the  surface 
to  be  dumped.  All  coal-mines  and  the  larger  metal-mines  em- 
ploy cars  as  the  common  carriers  for  the  mineral.  They  are 
loaded  as  stated  and  drawn  through  the  haulage- ways  to  the  foot 
of  the  shaft,  or  slope,  through  which  they  are  hoisted  to  day- 
light. 

Occasionally  when  the  pitch  of  the  slope  exceeds  40°  the  cars 
are  emptied  at  the  level  landing  into  skips,  which  are  hoisted  to 
the  surface,  while  the  car  returns  to  its  loading  station.  Slope 
carriages  are  also  used,  upon  which  the  car  is  delivered  and 
drawn  to  the  surface  to  be  emptied.  In  vertical  shafts  the  car 
is  hoisted  to  the  surface  on  a  cage,  and  returned  underground 
when  emptied. 

The  several  conveyances  named  will  be  here  discussed  together 
with  the  hoist-rope,  signalling  and  safety  devices,  and  the  method 
of  dumping  at  the  surface. 

Hoist-ropes. — Hemp,  aloe,  fibre,  and  iron  or  steel  wire  are 
used  for  hoisting  purposes  and  for  the  transmission  of  power  in 

196 


HOISTING  MACHINERY,  ETC.  197 

mines.  They  are  wound  into  strands  and  ropes  to  furnish  a 
strong,  flexible,  light  means  of  haulage.  The  hempen  ropes  for 
hoisting  purposes  have  been  supplanted  by  the  flexible  steel  wire 
for  many  reasons,  prominent  among  them  being  the  great  weight 
of  the  former  compared  with  its  strength  and  the  unreliability  of 
its  fibres.  So  frequently  are  the  hempen  ropes  built  up  of  strands 
of  discarded  ropes  that  long  service  is  never  obtained  from 
them.  This  fact,  together  with  their  absorption  of  moisture 
while  in  use,  contributes  to  their  rapid  decay. 

Steel-wire  Ropes. — The  successful  production  of  fine  iron  or 
steel  wire  and  its  conversion  into  flexible  ropes  for  hoisting  has 
led  to  the  abandonment  of  various  vegetable  fibres.  The  usual 
number  of  strands  in  a  multi-wire  rope  is  six,  with  nineteen 
wires  in  each  strand,  or  sometimes  only  seven  wires  are  used 
to  the  strand.  The  latter  are  more  frequently  employed  for 
guide-ropes  than  for  motor  purposes,  being  less  pliable  than 
the  former. 

The  increased  strength  imparted  to  metallic  wires  by  draw- 
ing, and  the  high  grade  of  steel  wire  as  compared  with  iron,  at 
once  recommend  it  for  any  purpose  where  strength  and  light- 
ness are  requisites.  Good  crucible-steel  wire  has  an  ultimate 
breaking  strength  of  75  to  100  tons  per  square  inch.  Tem- 
pered steel,  usually  known  by  the  indefinite  name  of  plow- 
steel,  is  a  high-grade  cast  steel,  used  for  ropes  of  the  greatest 
tensile  strength. 

Standard  Ropes. — Hoisting-ropes  in  the  United  States  are 
made  with  hemp  centres,  around  which  are  twisted  six  strands  of 
nineteen  wires  each,  the  individual  wires  being  about  one-fifteenth 
the  diameter  of  the  rope.  The  hemp  centre  imparts  flexibility 
and  furnishes  a  soft  yielding  core  upon  which,  with  a  minimum 
of  injury  or  friction,  the  strands  may  "creep"  as  the  rope  passes 
around  the  drum  and  sheave.  When  great  flexibility  is  re- 
quired, each  strand  has  also  its  own  centre,  but  this  is  unneces- 
sary for  ordinary  hoisting  service.  On  account  of  the  hardness 
and  diminished  flexibility  of  plow-steel  wire,  hoisting-ropes  of 
this  metal  are  often  made  with  eight  instead  of  six  strands,  the 


198  MANUAL  OF  MINING. 

wires  being  of  smaller  gauge  and  therefore  more  flexible.  Though 
the  enclosing  figure  of  the  hoisting- rope  is  a  polygon,  the  rope  is 
usually  called  a  round. 

The  Lay  of  the  Rope. — By  this  is  meant  the  twist  or  pitch  of 
the  wires  in  the  strands,  and  of  the  strands  in  the  rope.  As 
ordinarily  constructed,  the  lay  of  the  wires  is  opposite  to  the  lay 
of  the  strands.  The  former  is  as  long  as  possible  in  order  that 
the  fractional  wear  on  each  rope  shall  be  distributed  over  a  great 
length  rather  than  to  have  a  short  twisted  strand  on  which  the 
wear  is  concentrated.  The  so-called  Langlay  ropes  have  a 
longer  twist  to  each  wire,  thus  distributing  the  wear  over  a  greater 
'ength  and  giving  a  smoother  surface. 

Locked-coil  Ropes. — In  order  to  increase  the  smoothness  of 
the  rope  and  thus  to  reduce  wear,  ropes  are  frequently  made  of 
wires  which  are  locked  together  or  dovetailed.  The  locked- 
coiled  rope  is  of  this  type,  and  prevents  the  ends  of  the  broken 
wires  from  protruding.  Its  wires  are  built  of  a  specially  shaped 
cross-section  and  wound  in  concentric  layers,  the  successive 
layers  being  twisted  in  opposite  directions.  The  exterior  layer 
presents  a  perfectly  smooth  surface. 

Flat  Ropes  are  employed  with  winding  reels  where  it  is 
desirable  to  aid  the  engine  without  the  use  of  a  long  conical  drum. 
Moreover,  the  tendency  to  uncoil  which  exists  in  the  several 
strands  and  wires  of  round  rope  develops  a  whirling  tendency 
upon  the  cage  as  it  travels  in  the  shaft.  This  the  flat  rope  prevents. 
Flat  ropes  are  heavier  than  a  round  rope  of  the  same  strength, 
but  are  of  shorter  life,  costing  more  and  requiring  greater  care. 
A  flat  rope  is  composed  of  an  even  number  of  loosely  twisted  four- 
stranded  round  ropes,  without  hemp  centre,  laid  side  by  side  and 
sewed  through  the  wires.  The  lay  of  any  two  adjacent  strands  is 
in  opposite  directions.  The  sewing  wires,  from  eight  to  twelve  in 
number,  are  passed  from  side  to  side  through  the  centres  of  the 
strands.  Flat  ropes  range  in  width  and  thickness  from  2^'Xf" 
to  9"XJ". 

Although  the  breaking  strength  of  a  flat  rope  is  nearly  the 
same  as  that  of  a  round  rope  of  equal  weight  per  foot,  its  safe 


HOISTING  MACHINERY,  ETC.  1 99 

working  load  is  considerably  less,  due  to  the  practical  impossibility 
of  so  constructing  the  rope  that  all  the  strands  will  stretch  uni- 
formly and  be  equally  loaded  in  service. 

Taper  Ropes. — As  the  depth  of  shafts  increases,  the  use  of 
round  wire  rope  of  uniform  cross-section  is  attended  with  many 
difficulties.  The  diameter  of  the  rope  becomes  greater,  the  in- 
equality of  the  work  upon  the  engine  is  more  marked,  and  an 
excessively  large  of  drum  will  be  required.  There  is  a  limit  beyond 
which  a  rope  of  uniform  section  cannot  safely  carry  its  own  weight. 
At  12,000  feet  of  depth,  exclusive  of  any  live  load,  the  rope  has 
reached  the  limit  of  its  safe  working  stress.  The  additional  weight 
of  car  and  load  reduces  this  materially.  With  the  margin  provided 
for  safety,  the  maximum  depth  is  3000  feet.  A  given  rope  for 
carrying  men  having  a  factor  of  safety  of  15  when  less  than  1000 
feet  deep  has  a  factor  of  but  10  at  2500  feet  and  only  7  at  3000 
feet. 

Meanwhile  the  cross-section  of  the  rope  at  the  bottom  is  exces- 
sively heavy,  having  little  load  to  support.  If  the  cross-section 
be  such  as  will  support  the  live  load,  cage,  and  car,  then  the  cross- 
sections  of  the  rope  to  the  top  may  increase  in  area  by  an  amount 
equal  to  that  sufficient  to  support  the  additional  weight  of  rope 
below  that  point.  Hence  tapering  ropes  are  employed  in  special 
cases,  being  of  uniform  strength,  with  such  a  factor  of  safety  as  may 
be  desired.  Theoretically,  the  use  of  a  tapering  rope  increases 
the  limit  of  depth  which  can  be  reached  within  mining  possibili- 
ties, but  practically  it  is  not  a  popular  hoister.  The  round  tapering 
rope  is  made  by  discontinuing  at  intervals,  throughout  the  length, 
a  single  wire  at  a  time.  The  cross-section  of  the  rope  is  uni- 
formly decreased  toward  the  lower  end,  the  thickness  at  any 
point  being  sufficient  to  carry  safely  the  load  at  that  point. 
One  round  taper  rope  is  of  steel  of  five  sections,  varying  in 
diameter  from  i|  to  ^  inch.  The  sections  are  spliced  to- 
gether and  the  ends  of  the  wires  soft-soldered.  In  some  Eu- 
ropean districts  flat  taper  ropee  of  Manila  fibre  are  still  employed 
for  deep  shafts. 

The  taper  rope  is  not  well  adapted  to  long  slopes,  for  its  lower 


200 


MANUAL  OF  MINING. 


portion,  which  is  the  smaller  end,  traverses  a  greater  distance  than 
any  other  portion  of  the  rope  and  is  subjected  to  the  greatest  amount 
of  abrasion.  This  is  the  reverse  of  the  situation  in  a  shaft,  where 
the  thickest  part  of  the  rope  has  the  wear. 

The  Strength  of  a  Wire  Rope. — This  is  greater  than  that  of  a 
steel  rod  of  equal  cross-section  and  same  material,  because  of  the 
increased  strength  imparted  to  wires  during  the  process  of  draw- 
ing; but  it  is  not  equal  to  the  sum  of  the  strengths  of  the  indi- 
vidual wires. 

The  approximate  weight  of  a  wire  rope  in  pounds  per  foot  is 
ascertained  by  multiplying  the  square  of  its  diameter  in  inches 
by  1.58 

CAST -STEEL  HOISTING-ROPES: 

THEIR   ULTIMATE   STRENGTH,   MAXIMUM   SAFE   STRENGTH,   AND   WORKING   LOAD. 


6  Strands, 

6  Strands, 

Diameter 

Approxi- 
mate Cir- 

Esti- 
mated 

Proper 

79  Wires  each. 

1  9  Wires  each. 

in 
Inches. 

cumfer- 
ence in 

Weight 
per  Foot 

Working 
Load,  L. 

Ultimate 

Maximum 

Ultimate 

Maximum 

Inches. 

in  Lbs. 

Strength, 

Safe 

Strength, 

Safe 

Lbs. 

Stress,  5. 

Lbs. 

Stress,  S. 

if 

4* 

3.00 

116,000 

38,667 

124,000 

41,333 

J-i 

4 

2-45 

I 

96,000 

•3,2,000 

100,000 

33,333 

ij 

2.00 

"""!  co 

80,000 

26,666 

84,000 

28,000 

J 

3 

1.58 

%% 

64,000 

2I>333 

68,000 

22,000 

i 

$ 

2 
if 

I.  2O 
0.89 
O.62 

o.:;o 

0.39 

um  safe  stn 
bending  sti 

48,000 
37,200 
26,400 

21,200 
16,800 

16,000 
12,400 
8,800 
7,067 
5,600 

5?,CXDO 

38,800 
27,200 

22,000 
I7,600 

17,333; 
12,933 

7.335 

5,867 

$6 

1^ 

0.30 

J 

13,200 

4,400 

I3,600 

4,533 

• 

ji 

0.22 

a 

9,6OO 

3,200 

IO,OOO 

3,333 

%» 

! 

0.15 

6,800 

2,267 

6,800 

2,267 

* 

i 

O.  IO 

5,600 

1,867 

4,800 

1,600 

The  Minimum  Radius  of  Curvature. — The  minimum  drum 
diameter  admissible  for  ordinary  service  is  from  50  to  TOO  times 
the  diameter  of  the  rope.  By  using  the  larger  ratio  the  average 
life  of  hoisting-ropes  would  be  materially  lengthened.  The 
English  make  the  sheave  diameter  a  multiple  of  X2.  The  length 
of  the  arc  of  contact  between  rope  and  sheave  and  the  angle 
between  the  two  branches  of  rope  are  not  the  same  as  would  be 


HOISTING  MACHINERY,  ETC. 


the  case  with  a  belt  or  Manila  rope.  A  large  part  of  the  total 
strength  of  the  rope  is  consumed  in  the  mere  strain  of  bending, 
and  this  is  wholly  a  question  of  radius  of  curvature  depending 
upon  the  diameter  of  the  sheave. 

This  bending  resistance  increases  the  tension  on  the  wires 
of  the  ropes  and  reduces  the  available  load  which  it  can  safely 
carry  by  that  amount.  The  continual  bending  naturally  reduces 
the  life  of  the  rope. 

In  the  following  table  are  given  the  values  of  the  radii  of 
curvature,  R,  in  inches,  assumed  by  new  steel  ropes  of  stated 
diameters  having  a  tension  of  one  pound  upon  them  and  having 
an  angle  of  deflection  or  of  bending  of  i°. 

THE  RADIUS  OF  CURVATURE  R  OF  STEEL-WIRE  ROPE,  IN  INCHES,  HAVING 
i°  DEFLECTION  AND  i  LB.  TENSION. 


Diameter. 

7-wire. 

iQ-wire. 

1 

257,270 

84,210 

^ 

592,97° 

220,950 

I 

1,241,510 

454,120 

I 

2,315,030 

860,400 

I 

3,97^290 

1,430,850 

4 

6,386,740 

2,246,676 

i| 

10,049,650 

3,369,820 

To  obta'n  the  radius  of  curvature  under  similar  conditions 
for  any  other  angle  of  deflection,  it  is  only  necessary  to  divide  the 
corresponding  quantity  by  the  number  of  degrees  of  bend.  Thus 
a  new  i-inch  ip-wire  rope,  for  a  tension  of  one  pound  upon 
it,  would  have  a  curvature  whose  radius  is  1,430,850  inches 
when  bent  through  an  angle  of  i°.  If,  however,  it  is  bent  90° 
over  a  given  sheave,  the  radius  of  the  curve  which  it  will  assume 
then  becomes  15,898  inches.  When  the  tension  becomes  1000 
Ibs.  the  curvature  which  it  will  assume  will  have  a  radius  of 
16  inches. 

Bending  Rope  Around  Curves. — When  a  rope  is  carried  over 
a  sheave  with  an  angle  of  contact  or  wrap  exceeding  30°,  the 
diameter  of  the  sheave  must  be  of  liberal  dimensions  When, 
however,  the  rope  is  slightly  deflected  by  a  pulley  or  sheave, 


202  MANUAL  OF  MINING. 

the  stress  put  upon  the  rope  in  bending  is  small.  When  it  is 
bent  over  a  succession  of  sheaves  at  close  intervals,  as  in  carry- 
ing a  rope  around  curves  in  a  roadway,  the  aggregate  bending 
stress  may  become  significant.  The  rope  assumes  between  each 
pair  of  sheaves  a  curvature  depending  upon  the  tension  it  carries 
and  the  amount  of  each  deflection. 


FlG.  75.  —  Turning  the  Haulage-rope  at  Curves. 

The  distance  between  the  sheaves  may  be  determined  from 
the  formulae  below  when  the  tension  is  known  and  the  radius  of 
curvature  has  been  ascertained  by  the  preceding  table. 

Let  R  =  radius  of  curvature  i°  bend  with  i  Ib.  tension  for  rope 

of  diameter  X  (from  the  preceding  table); 
r=  radius  of  curvature  assumed  by  rope  under  tension  T\ 
d=  diameter  of  individual  wires  in  rope,  inches; 
w=  number  of  wires  in  rope; 
T=  tension  on  the  rope,  Ibs.; 
s=  distance  between  sheave  centres,  inches; 
a  =  angle  of  deflection  over  a  pulley,  and 
p=  radius  of  centre  line  of  sheaves,  inches. 
Then 


The   sheaves  should  be   laid  on   a  curve  of  radius  r  —  that 


HOISTING  MACHINERY,  ETC.  203 

is,  r=p't  d=o.inX  for  a  y-wire  rope  and  0.066  for  a 
rope. 

R 


s  =2p  sin  %a. 

The  Influence  of  Bends  on  the  Durability  of  a  Rope.  —  The 

diameter  of  the  sheave  over  which  the  rope  is  bent  must  in  all 
cases  exceed  the  minimum  diameter  of  curvature  assumed  by 
the  rope  under  the  conditions  as  calculated  above.  Not  only 
is  the  life  of  the  rope  increased,  but  also  its  bending  stress  is  con- 
siderably diminished  and  the  safe  working  load  of  the  rope 
correspondingly  increased.  Under  ordinary  conditions  of  wear, 
the  life  of  a  rope  for  mining  purposes  averages  about  seventeen 
months,  after  which  time  the  abrasion  of  some  wires  and  the 
breaking  of  others  render  it  unsafe.  This  assumes  proper  care 
in  lubrication,  in  testing,  care  during  hoisting  to  avert  shocks, 
and  the  use  of  liberal-sized  sheaves.  Tests  upon  wires  of  various 
sizes  passing  over  pulleys  of  various  dimensions  have  demon- 
strated that  a  wire  subjected  to  repeated  bending  soon  breaks, 
the  time  of  rupture  being  hastened  as  the  diameter  of  the  sheave 
or  the  angle  of  the  bend  is  small.  For  example,  a  wire  of  No.  20 
B.W.G.,  subjected  to  repeated  bending  over  a  5  -inch  pulley, 
broke  after  making  i5j2oo  turns;  one  of  the  same  size  and  pre- 
sumable strength  made  453,000  turns  over  a  pulley  of  24  inches 
in  diameter  before  breaking. 

EXAMPLE.  —  It  is  desired  to  carry  a  ii-inch  ip-wire  steel  rope  around  a  curve 
(Fig.  75)  whose  track  radius  is  50  feet,  by  guide-sheaves  44  inches  inside 
of  the  track  centre.  What  should  be  the  distance  between  the  sheaves? 

T=66o,    X=i.25,    r=  1,430,900  inches,    and    p=  554  inches.    Then 
a  =3°  oo'    and    5=2  feet  5  inches. 

The  Working  Load.  —  This  is  a  fraction  of  the  ultimate  strength 
of  the  rope,  depending  upon  the  margin  of  safety  desired.  For 
ordinary  conditions  of  hoisting  the  factor  of  safety  is  taken  at  7  for 
hoisting-ropes  and  6  for  those  employed  on  slopes.  The  maximum 
stress  to  which  the  rope  can  be  subjected  should  be  less  than 


204 


MANUAL   OF  MINING. 


PLIABLE  HOISTING  WIRE  ROPE 

6  STRANDS- 19  WIRES  EACrf-CAST  STEEL 

DENVER  ENGINEERING  WORKS 

DECEMBER   1,  1903. 


NT*  NT  i*i      100 

FIG.  76.— The  Net  Working  Load  of  Ropes  over  Sheaves. 


HOISTING  MACHINERY,  ETC.  205 

one-seventh  or  one-sixth,  respectively,  of  the  ultimate  strength 
of  the  rope.  A  larger  factor  of  safety  would  unnecessarily  in- 
crease the  size  of  the  rope,  which  in  turn  increases  the  dead  load, 
due  to  its  own  weight  and  the  diameter  of  the  sheaves  over  which 
it  must  pass. 

The  stresses  to  which  a  rope  is  subjected,  in  addition  to  the 
live  load  to  be  carried,  are  those  due  to  its  own  weight  and  friction, 
its  bending  over  the  sheave,  its  rigidity,  and  the  shock  induced 
at  the  time  of  starting.  Some  of  these  can  be  calculated,  but 
the  tension  induced  by  the  rigidity  of  the  rope  is  not  so  easily 
ascertained.  An  initial  strain  is  set  up  in  the  wires  by  the  bend- 
ing around  the  drum  or  sheave  which  is  not  equally  distributed 
throughout  the  rope.  Its  value  depends  upon  the  sharpness  of 
the  bend  and  is  ascertained  by  the  formula  given  below.  A 
small  sheave  and  a  large  angle  of  deflection  produce  a  large 
bending  stress. 

The  Resistance  Due  to  Bending.  —  This  must  be  determined 
before  the  available  load  which  the  rope  can  carry  can  be  ascer- 
tained. Usually  the  live  load  represents  the  difference  between  the 
safe  or  minimum  allowable  tensile  stress  and  the  stress  due  to 
bending.  In  deep  shafts  regard  must  also  be  taken  for  the 
weight  of  the  rope,  which  is  to  be  subtracted  from  the  working 
load. 

The  bending  stresses,  k,  which  are  developed  in  a  rope  when 
bending  over  a  sheave  or  hoisting-drum  through  an  angle  ex- 
ceeding 30°  are  indicated  below. 

Let  X  =  the  diameter  of  the  steel  rope  in  inches; 
r=the  radius  of  the  sheave  or  drum  in  inches; 
/=the  thickness  of  the  flat  wire  rope  in  inches; 
k  =  the  bending  stress  in  pounds. 

Then  for  a  y-wire  rope 


r+o.3X  ' 
for  a  ip-wire  rope 

367,cooX3 
r+o.$X  '. 


206  MANUAL  OF  MINING. 

for  a  flat  wire  rope 

89,465* 
r+2.23/' 

The  diagram  in  Fig.  76  graphically  shows  the  net  working 
load  which  a  given  rope  of  pliable  steel  can  carry  over  a  pulley  of 
a  given  diameter.  This  diagram  is  based  on  both  formulEe,  and 
is  taken  from  a  treatise  of  the  Trenton  Iron  Co.  on  Wire-rope 
Transmission.  As  an  illustration  of  its  use  let  it  be  required  to 
determine  the  net  load  for  a  i-inch  rope  on  a  ico-inch  sheave.  At 
the  intersection  of  the  vertical  line  from  the  point  on  the  top  line 
marked  100  inches  to  the  curve  representing  the  i-inch  rope,  the 
horizontal  line  carried  to  the  right-hand  edge  shows  15,500  Ibs. 
as  the  net  load.  It  will  be  noticed  that  a  2-inch  rope  on  a  76- 
inch  sheave  has  a  net  working  load  not  greater  than  that  of  a 
f-inch  rope  on  the  same  sheave.  This  diagram,  therefore,  indi- 
cates the  reduction  in  the  capacity  of  the  rope  due  to  bending 
stress. 

Elastic  Connections  between  Rope  and  Cage. — Usually  while 
lowering  to  a  platform  an  excess  of  rope  is  paid  out  and  tends  to 
coil  upon  the  top  of  the  cage.  This  operates  disastrously  upon 
the  wires  of  the  rope,  which  become  bent  and  ultimately 
broken.  A  flexible  connection,  therefore,  is  provided  of 
a  length  sufficient  to  allow  for  the  probable  slack.  This 
consists  of  a  short  length  of  chain  with  a  snap- hook  (Fig. 
77),  which  can  be  released  when  desired.  Owing  to  the 
shock  which  these  links  receive,  they  must  be  carefully 
examined  at  short  intervals.  In  some  mines  it  is  the 
practice  also  to  anneal  the  iron  by  heating  to  a  red  heat 
and  cooling  slowly.  This  is  done  once  a  month. 

These  connections  should  be  as  short  as  practicable 
in  order  to  diminish  the  amount  of  shock  induced  at 
starting.      Carelessness   in   paying  out   excessive   slack 
'  reduces  the  life  of  the  rope  and  increases  the  stress  at 
starting. 

Results  of  observations  in  Prussian  mining  districts  during 


HOISTING  MACHINERY,  ETC  207 

the  past  twenty  years  showed  the  effect  due  to  repeated  shocks 
at  starting  to  be  more  disastrous  to  the  flat  than  to  the  round 
steel  ropes.  In  certain  districts  6  per  cent  of  flat  steel  ropes 
broke  suddenly  out  of  1419  in  use;  12  per  cent  of  flat  charcoal- 
iron  ropes  gave  way  suddenly;  7  per  cent  out  of  124  flat  Manila 
ropes  broke  without  previous  warning;  2  per  cent  of  5527  round 
steel  ropes  gave  way,  and  10  per  cent  of  1350  round  charcoal 
ropes  failed  without  giving  previous  warning. 

The  tension  due  to  jerks  is  twice  that  due  to  the  dead  load 
with  6  inches  of  slack,  and  more  if  the  slack  is  greater. 
Let  T=the  total  tension  at  starting; 

a= acceleration  at  beginning  of  hoist; 

h= slack  in  inches; 

b  =  elongation  of  rope  due  to  total  load,  W. 

W= weight  of  car,  cage,  mineral,  and  rope. 


Rope  Sockets. — There  are  two  types  of  sockets  for  round  ropes 
— the  conical  and  the  double-pin,  the  former  being  the  stronger. 
The  conical  socket  (Fig.  78)  is  slipped  on  to  the  rope,  the  wires 
are  untwisted,  hemp  centres  cut  out,  the  wires  bent  back  and 
forth  into  a  tangled  mat  to  fill,  as  nearly  as  possible,  the  conical 
socket,  which  is  then  slipped  into  place.  This  is  slightly 
heated,  and  soft  lead  poured  in  to  solidify  the  mass. 
The  socket  and  rope  are  surrounded  with  wet  clay  to 
prevent  heating  of  the  wires  beyond.  The  double-pin 
is  treated  in  the  same  manner,  but  its  connection  with 
the  chain  is  by  a  pair  of  pins  through  the  links,  instead 
of  a  ring  for  hooking,  as  in  the  former  case.  A  "goose- 
neck" socket  consists  of  a  pair  of  trough- shaped  tongs, 
bent  to  a  loop,  and  riveted  to  the  rope  by  three  or 
four  rivets,  driven  cold.  Flat  ropes  have  riveted  to 
them  shackles  with  eyes,  which  receive  the  first  link  g 
of  the  chain.  Six  inches  of  the  end  are  untwisted  and 
doubled  back,  bound  with  wire,  the  shackles  slipped  on,  riveted 


208 


MANUAL  OF  MINING. 


?.    ?.     8 


HOISTING  MACHINERY,  ETC.  209 

through  the  rope,  and  the  hoops  finally  slipped  on  and  driven 
tight. 

In  securing  the  rope  on  the  drum  it  is  only  necessary  to  con- 
tinue several  extra  coils  of  the  rope,  insert  the  end  through  the 
wooden  lagging,  and  fasten  it  on  the  hub  or  shaft;  or,  instead, 
the  end  may  be  bolted  to  me  arm  of  the  casting.  If  the  fastening 
have  but  a  lo-lb.  grip  on  the  rope,  it  will  resist  a  weight  of  90 
Ibs.  if  there  is  only  one  coil  around  the  drum;  if  there  are  two 
extra  coils,  800  Ibs.  will  not  budge  the  zo-lb.  grip;  with  three 
extra  coils  it  requires  7300  Ibs.;  while  with  four  it  has  a  65,000- 
Ib.  resistance  and  can  support  that  amount  of  tension. 

Wire  rope  is  spliced  in  the  same  manner  as  hemp.  The 
strands  are  unlaid  for  3  feet,  and  each  passed  over  one  and  under 
another  of  its  corresponding  strands  on  the  opposite  rope,  for 
a  like  distance;  the  free  ends  are  then  trimmed  off  close.  Short- 
twisted  rope  is  more  easily  spliced  than  one  of  long  twist. 

The  Head-frame  is  designed  to  withstand  the  stresses  from 
the  weight  of  the  dead  and  live  loads  on  the  rope  and  their  fric- 
tional  resistance  operating  vertically,  and  from  a  pull  of  the  engine 
along  the  inclined  rope.  The  latter  .exceeds  the  former  by  an 
amount  equal  to  the  frictional  resistance  at  the  sheave  and  its 
bending  stress.  The  frictional  resistance  may  be  taken  as  one  per 
cent  of  the  gross  load  on  the  journals,  and  the  bending  stresses 
•of  the  rope  may  be  determined  from  the  preceding  formulae. 
The  combination  of  these  two  forces  produces  a  resultant  operating 
in  a  direction  which  nearly  bisects  the  angle  between  the  two 
ropes.  From  these,  with  the  load,  may  be  obtained  the  direction 
of  the  resultant.  Its  line  should  fall  nearly  central  within  the 
faase  of  the  head-frame,  when  the  latter  is  of  wood.  It  is  not  so 
essential  if  the  derrick  is  of  steel. 

The  frame  over  the  shaft  is  built  in  one  of  two  patterns.  The 
form  of  the  derrick  is  essentially  two  vertical  right- angle  triangles, 
each  upright  and  brace  being  respectively  parallel  to  the  two 
directions  of  the  rope.  The  apex  of  the  triangle  should  be  at  the 
centre  of  the  axis  of  the  sheave.  The  triangular  frames  are 
connected  in  a  transverse  direction  by  braces  and  ties.  Instead 


2io  MANUAL   OF  MINING. 

of  single  sticks  for  the  uprights,  a  vertical  frame  is  built  of  four 
uprights  surrounding  the  shaft- opening  with  two  or  three  struts 
projecting  from  their  top  toward  the  hoisting-engine  and  inclined 
to  furnish  ample  base  for  the  derrick. 

In  the  other  pattern  the  four  posts  are  batired  outward  to 
afford  an  ample  square  base  of  support  without  the  use  of  any 
additional  struts  (Fig.  80).  The  frames  may  be  built  of  wood  or 
steel,  the  latter  being  extensively  employed  in  collieries  where 
the  speed  of  hoist  is  great  (Fig.  81).  The  steel  frames  are 
built  by  construction  companies  to  order  and  are  composed 
of  the  usual  structural  shapes  in  steel. 

The  foundation  may  be  either  masonry  pillars  or  concrete, 
extending  some  distance  outside  of  the  lines  of  the  shaft.  The 
lower  base  consists  of  three  longitudinal  sills  framed  with  three 
cross-sills.  Bolted  to  the  former  are  cast-iron  shoes  to  receive 
the  uprights  and  braces.  The  base-frame  is  bolted  and  anchored 
to  heavy  timbers  buried  in  the  ground  or  in  the  concrete. 

The  Height  of  the  Derrick. — The  essential  requirements  of  the 
head-frame  are  stability  and  safety.  The  former  is  obtained  by 
a  broad  base  and  a  rigid  construction  of  frame;  the  latter  is 
obtained  by  its  construction  of  fire-proof  metal  and  a  height 
sufficient  to  satisfy  the  conditions  of  hoisting.  The  risk  of  fire, 
the  effect  of  weather,  the  working  of  the  joints,  and  the  difficulty 
of  securing  sound  sticks  of  requisite  length  render  the  adoption  of 
material  other  than  wood  advisable.  Height,  which  is  an  essen- 
tial feature  of  the  head-frame,  is  difficult  to  attain  without  a 
rigid  frame,  and  this  implies  steel  construction.  This  is  the 
rule  at  large  mines.  Steel  frames  are  more  desirable  and  immeas- 
urably more  rigid  than  wooden  ones.  To  withstand  injury  from 
the  continual  vibration,  pin  connections  are  more  suitable  than 
riveted  connections. 

The  height  of  the  derrick  should  be  such  as  will  furnish 
security  against  overwinding  by  providing  sufficient  margin  within 
which  the  engineer  may  stop  the  hoist.  The  greater  the  speed 
of  the  engine  the  greater  is  the  necessity  for  a  high  derrick.  The 
minimum  allowance  is  that  which  corresponds  to  the  length  of 


HOISTING  MACHINERY,  ETC.  211 

two  coils  of  rope  on  the  drum  plus  the  height  of  the  cage.  The 
distance  from  the  landing  station  for  cars  to  the  sheave  would 
give  the  engineer  four  or  five  seconds  in  which  to  secure  control 
of  his  engine.  Usually  the  derrick  is  of  a  height  twice  the  amount 
stated.  In  anthracite  regions  it  is  often  more  than  this. 

Steel  frames  in  some  instances  are  65  feet  in  height,  the  base 
of  the  first-mentioned  type  being  about  16  feet  square  with  the 
sides  extended  for  the  braces.  Of  the  second  type  there  are 
numerous  steel  frames  with  base  of  25  to  30  feet  square,  with  a 
height  of  50  feet.  Derricks  are  frequently  extended  to'  a  height 
above  that  of  the  tipple  or  breaker  platform  to  permit  of  a  con- 
tinual descent  of  the  mineral  from  the  time  of  unloading  to  the  time 
of  shipment.  In  such  cases  as  these  the  head-frame  is  over  100 
feet  in  height.  Those  in  the  zinc-mining  districts  of  Missouri 
.are  of  the  four-post  type,  30  feet  high,  with  suitable  braces.  Some 
in  the  anthracite  district  are  as  high  as  130  feet. 

The  hoist-frame  is  rarely  enclosed,  and  preferably  is  not 
housed  because  of  the  danger  from  fire.  A  fire  breaking  out  in 
the  building  over  a  shaft  would  be  uncontrollable  with  a  draft 
from  below,  besides  risking  the  communication  of  the  fire  to 
timbers  of  the  shaft.  A  fire  communicated  to  the  hoist-frame 
from  below  would  endanger  the  only  means  of  exit  which  the 
men  have  from  underground. 

The  tipple  type  of  head-frame  is  built  of  bents,  suitable  for  the 
screens  and  pockets.  It  connects  with  and  extends  from  the 
head-frame  in  a  direction  opposite  to  that  of  the  back-stays.  It  is 
of  wood  or  steel  with  all  members  carefully  calculated.  The 
rhythmic  motion  set  up  by  the  vibrating  screens  must  be  pro- 
vided for,  and  the  design  varies  radically  between  one  having  a 
dump  in  the  plane  of  the  hoist-rope  and  one  in  which  the  shocks 
of  dumping  are  at  right  angles  to  that  plane. 

Slopes  require  no  high  head-frame,  for  the  line  of  track  extends 
directly  from  the  shaft  mouth  to  the  elevated  structure,  and  in 
reality  forms  the  back- stay  for  resisting  the  overturning  tendency 
of  the  structure.  This  is  the  strongest  possible  type  of  frame, 
for  the  resultant  falls  inside  of  the  base  of  the  tipple. 


212 


MANUAL  OF  MINING. 


HOISTING  MACHINERY,  ETC. 


213 


The  Sheave. — Upon  the  top  of  the  uprights  is  a  frame  which 
supports  the  sheave  on  a  short  horizontal  axle.  It  may  be  built 
directly  upon  the  braces  and  uprights  of  the  head-frame  or  upon 
the  platform  above  it.  The  dimensions  of  the  sheave  are  deter- 
mined by  the  size  of  the  rope.  It  consists  essentially  of  a  spider 
frame  carrying  a  grooved  circumference  filled  with  some  material 


FIG.  81. — A  Steel  Head-frame  at  a  Bituminous  Colliery. 

which  has  a  large  coefficient  of  friction,  as,  for  instance,  wooden 
blocks  on  the  end,  or  rubber.  These  latter  prevent  slip  and 
render  the  hoisting  more  secure,  as  well  as  protect  the  rope  from 
excessive  wear.  Its  hubs  are  double,  connected  to  the  rim  by 
rods  let  into  sockets.  The  entire  construction  is  made  as  light  as 
possible  consistent  with  strength,  to  reduce  the  inertia  and  avoid 
the  abrasion  that  occurs  with  a  heavy  wheel  continuing  to  revolve 
after  hoisting  has  ceased. 


a  14  MANUAL  OF  MINING. 

The  cost  of  the  high  steel  head- frames  of  a  mine  averages 
$32.20  per  ton  of  material,  including  the  design  and  details.  The 
erection  averages  $13.00  per  ton  when  the  connections  are  riveted. 

As  regards  the  cost  of  wooden  head-frames,  the  following  may 
afford  a  suggestion  for  average  conditions.  The  four- post  frame 
for  small  mines  of  31  feet  requires  1800  feet  of  board  measure 
and  costs  from  $100  to  $150  for  labor  and  supplies.  Larger 
frames  without  back- stays  for  the  height  of  50  feet  consume 
14,000  feet  of  lumber  and  cost  $550.  Tipple  head-frames  for 
vertical  shafts,  with  60  feet  of  height,  require  15,000  feet  of  board 
measure  and  cost  about  $800.  Slope  tipples  for  inclined  shafts, 
built  to  a  height  of  31  feet,  are  completed  for  $3300,  using  40,000 
feet  of  board  measure. 

Cage  Indicators. — Several  independent  devices  are  used  to 
indicate  the  position  of  the  cage  in  the  shaft  at  all  times  during 
its  hoist.  Automatic  indicators  are  installed  in  view  of  the 
hoisting  engineer.  An  indicator  has  a  pointer  which  moves 
around  a  dial  or  vertically  along  the  side  of  a  column,  the  dial 
or  the  column  being  graduated  to  the  depths  of  the  shaft-  or 
slope-landings.  A  string  winding  on  a  miniature  counterpart 
of  the  hoisting-drum  moves  the  index  a  distance  commensurate 
with  that  of  the  cage  and  shows  the  latter's  position  after  starting 
and  its  proximity  to  the  landing  station  during  the  hoist.  The 
approach  of  the  cage  is  also  indicated  by  an  electric  communi- 
cation automatically  established  by  the  cage  in  its  ascent.  In 
this  arrangement  the  cage  makes  an  electric  connection  as  it 
passes  a  given  point  in  the  shaft  below  the  point  of  landing, 
which  is  signalled  to  the  engineer.  A  second  notice  is  given  in 
a  similar  manner  when  the  cage  is  just  about  to  reach  the  land- 
ing. This  notification  is  independent  of  the  automatic  devices 
for  the  guidance  of  the  engineer  in  braking  the  engine.  In  addi- 
tion to  these  means,  the  brakeman  or  landing-man  at  the  head 
of  the  shaft  having  charge  of  the  cars  may.  signal  the  engineer 
as  the  cage  approaches  the  top  and  furnish  another  control, 
which,  being  sentient,  is  always  to  be  preferred  to  any  mechanical 
device,  no  matter  how  perfect. 


HOISTING  MACHINERY,  ETC.  215 

• 

Signalling. — The  communication  between  the  employees  below 
and  the  engineer  above  may  be  had  by  the  use  of  a  wire  with  a 
bell  or  some  electric  annunciator,  or  by  a  speaking-tube  or  tele- 
phone. Whatever  may  be  the  means  employed,  it  is  essential 
that  it  should  be  quick  and  accurate,  presenting  no  opportunity 
for  mistakes.  The  early  clumsy  arrangements  depending  upon 
the  signal  of  the  gong,  bell,  or  triangle  which  is  struck  by  a  lever, 
operated  from  below  by  a  rope  or  wire,  are  untrustworthy.  The 
pull  on  the  rope  may  be  too  light  to  strike  the  blow,  or  more  strokes 
may  be  rung  up  than  are  intended,  and  there  is  no  means  of  making 
a  correction  before  the  engineer  will  have  proceeded  to  answer 
the  call.  Accident  may  then  ensue.  The  annunciator  is  ano'her 
method  which  is  quick  and  convenient  for  signalling  to  the  engineer 
or  between  stations,  but  its  scope  is  limited  to  a  very  small  set 
of  signals.  Unquestionably  electricity  offers  the  best  means 
of  signalling  between  stations  in  dry  shafts.  One  wire  or  a  couple 
of  wires  may  be  placed  within  easy  reach  of  those  in  the  cage, 
furnishing  a  means  for  its  occupants  to  signal  to  the  surface 
or  elsewhere  while  in  motion.  Such  wires  would  also  be 
convenient  for  those  engaged  in  examining  or  repairing  the 
shaft,  being  connected  to  electric  gongs  located  at  each  landing- 
station. 

A  uniform  code  of  signals  among  mining  men  is  an  eminently 
desirable  feature.  A  hoisting  engineer  at  a  new  mine  may  err 
with  serious  results  by  an  interpretation  of  a  signal  which  else- 
where represents  some  other  operation. 

The  Telephone. — There  need  be  no  argument  to  convince 
the  reader  of  the  advantages  of  the  speaking-tube  or  telephone. 
This  safe  natural  means  is  largely  employed.  It  is  rapidly  sup- 
planting the  electro-mechanical  gongs.  The  expense  of  laying 
out  a  mine  telephone  system  is  generally  much  less  than  that 
of  a  surface  system,  since  in  most  cases  the  wires  can  be  sup- 
ported by  means  of  porcelain  or  glass  insulators,  either  from 
the  roof  or  from  any  timber  which  is  used  on  the  sides  of  the 
road.  Their  greater  expense  is  overbalanced  by  the  great 
saving  in  time,  and  their  efficiency  in  case  of  accident.  For 


216  MANUAL   OF  MINING. 

deep  shafts,  telephones  are  almost  imperative;    in  fact,  mining 
legislation  in  some  States  compels  their  use. 

Bridging  telephones,  Fig.  82,  connected  to  the  metallic  sig- 
nal-line, used  in  connection  with  rope  haulage,  reduce  the  cost 
of  installation  to  that  of  first  cost  of  the  telephones  and  the  nomi- 
nal cost  of  connecting  up.  For  insulation,  rubber-covered  wire 


FIG.  82. — A  Mine  Telephone. 

will  serve  generally  in  localities  where  moisture  or  contact  with 
timber  might  cause  an  interruption  of  service.  When  an  in- 
stallation of  mine  telephones  will  warrant  the  extra  cost,  lead- 
covered  cables  should  be  employed. 

Means  of  Increasing  the  Safety  of  Ascent  and  Descent  into 
the  Mines. — The  necessity  for  the  introduction  of  safety  appliances 
of  this  kind  needs  no  argument  when  one  considers  that  the  engi- 
neer, from  a  point  somewhat  remote  from  the  elevator,  is  raising 
the  latter,  at  an  hourly  rate  of  speed  reaching  30  miles,  and  the 
load,  whose  weight  may  be  nearly  one  fourth  of  the  strength  of 
the  rope.  This  is  to  be  landed  promptly,  accurately,  and  without 
shock  at  the  precise  point.  In  this  work,  which  is  repeated  scores 
of  times  each  hour  and  with  an  engine  of  several  hundred  horse- 
power, with  no  guide  except  the  index  moving  about  a  small  circle, 
any  defect  in  the  operation  of  the  indicator,  or  any  momentary 
distraction  of  the  engineer's  attention,  may  result  disastrously 
and  the  engine  raise  the  cage  a  few  feet  beyond  the  point  in- 
tended, and  indeed  may  raise  it  over  the  sheave  at  the  top  of 
the  frame.  Again,  some  defect  in  the  operation  of  the  throttle 
or  of  the  brake  may  prevent  the  engineer  from  bringing  the  engine 


HOISTING  MACHINERY,  ETC.  217 

to  rest  as  promptly  as  desired,  and  one  revolution  or  two,  which 
may  wind  30  to  50  feet  of  rope,  may  bring  the  cage  to  the  danger 
limit.  Recalling  that  these  operations  are  performed  50,000 
times  per  year,  the  mental  strain  and  the  anxiety  of  the  engineer 
where  no  overwinding  device  is  provided  is  unquestionably 
serious.  Again,  coal  and  men  are  hauled  on  the  same  vehicle; 
and  if  no  proper  method  of  signalling  exists,  the  engineer  may 
hoist  the  latter  at  the  speed  and  to  the  place  of  dumping  the 
former. 

Though  the  number  of  accidents  arising  from  overwinding 
is  small,  nevertheless  it  is  sufficient  in  amount  to  warrant  the 
introduction  of  some  device.  There  are  two  general  classes  of 
these.  The  first  type  is  an  ample  detaching-hook,  and  the  second 
an  automatic  control  for  the  engine. 

Detaching-hooks. — Safety  contrivances  are  installed  at  the 
head-frame  for  preventing  a  cage  from  falling  when  severed 
from  the  rope  through  overwinding.  Their  design  contem- 
plates the  release  of  a  link  between  the  rope  and  its  cage,  and 
the  simultaneous  action  of  some  support  which  take?  hold  of  a 
portion  of  the  framework  of  the  headgear.  These  give  effective 
service  and  are  employed  very  extensively  in  Europe,  but  to  a 
lesser  extent  in  America. 

Ormerod's  hook  is  one  of  the  best  of  these  safety  links.  The 
apparatus  when  in  ordinary  use,  as  in  2,  Fig.  84,  is  wider  at  the 
bottom  than  at  the  top.  When  overwinding  occurs  the  link  is 
drawn  into  the  bell-mouthed  cylinder  just  below  the  sheave. 
This  lower  part  comes  into  contact  with  the  cylinder,  thereby 
closing  the  bottom  part  of  he  link  and  expanding  the  top  part. 
Its  projections  catch  over  the  top  of  the  cylinder,  while  at  the 
same  time  the  rope  shackle  A  is  forced  out  of  its  seat,  thus  being 
allowed  to  go  free;  the  bottom  shackle,  B,  drops  into  a  slot  and 
locks  the  link  firmly  in  its  position.  The  cage,  being  suspended 
from  the  chain,  cannot  fall  back.  To  prevent  the  possibility  of 
the  link  becoming  disarranged  in  ordinary  work,  a  small  pin  is 
inserted  centrally  through  the  plates,  which  pin  is  sheared  off  as 
the  apparatus  passes  into  the  cylinder. 


218 


MANUAL   OF  MINING. 


Middleton's  hook,  3,  Fig.  84,  operates  on  the  same  principle. 

King's  hook,  one  form  of  detaching-hook,  is  illustrated  in 
Fig.  83.  It  consists  of  two  outside  plates,  enclosing  two  inner 
ones,  capable  of  oscillating  about  a  strong  pin  which  passes 
through  the  series.  When  hoisting  has  reached  the  danger- 


FiG.  83. — King's  Detaching-hook. 


point,  the  wings  on  the  sides  of  the  latter  pair  strike  the  plates, 
the  jaws  permitting  the  rope  to  be  hoisted  over  the  sheave  and 
the  cage  to  fall.  At  the  same  time,  however,  the  projecting 
wings  on  the  hook  catch  on  the  protruding  edges  of  the  platform 
and  hold  the  cage. 

Walker's  safety  attaching-hook,  4,  Fig.  84,  has  a  loop  which 
encircles  the  hook  and  is  bound  by  it  to  two  copper  rivets.  These 
are  sheaved  when  the  hook  is  down.  The  jaws  then  open  and 
release  the  rope,  locking  the  suspension  jaws  on  the  disengaging- 
plate  and  holding  the  cage. 

Engine  Controllers. — This  class  is  a  more  or  less  flexible  con- 
nection between  the  shaft  and  the  engine  by  which  the  rage 


HOISTING   MACHINERY,  ETC 


^220  MANUAL   OF  MINING. 

-automatically  shuts  off  the  supply  of  steam  or  applies  the  brake 
when  it  reaches  a  certain  point. 

The  Visor,  5,  Fig.  84,  consists  of  a  governor  suitably  driven 
by  a  crank-shaft,  which,  if  the  speed  of  the  engine  exceeds  a  cer- 
tain point  at  the  end  of  the  hoist,  applies  the  foot-brake  through 
a  combination  of  levers,  and  closes  the  steam- valve. 

Buckets — Kibbles. — The  simplest  conveniences  for  mineral 
Jioist  are  employed  during  sinking  and  during  the  mining  opera- 
tions of  small  mines.  These  are  sheet-steel  tubs,  or  buckets, 
fitted  at  the  top  with  a  bale  of  round  iron  and  at  the  bottom  with 
a  small  ring.  The  bale  is  hooked  into  eyes  opening  into  a  strap 
which  is  bolted  to  the  sides  and  under  the  bottom  of  the  tubs. 
The  snap-hook  (Fig.  85)  on  the  hoisting-rope  catches  into  the 
ring  of  the  bale.  When,  however,  buckets  are  to  be  frequently 


FlG.  85.— Snap-hook.  FIG.  86.— Serpentine  Hook. 

cletached  from  the  rope,  serpentine  hooks  (Fig.  86)  are  used 
instead.  They  are  plain  cylinders  of  sheet  steel  or  boiler-iron, 
18  to  33  inches  in  diameter,  30  to  54  inches  deep,  having 
a  capacity  of  600  to  3000  Ibs.,  and  a  weight  of  about  one 
•quarter  of  their  contents.  Frequently  they  are  bellied  and  used 
In  slopes  on  skids.  The  buckets  are  loaded  from  chutes  at  the 
mill-holes  in  the  mine  (Fig.  87),  and  transported  on  the  trucks 
to  the  shaft  and  raised  to  the  surface. 

The  dumping  of  the  bucket  at  the  surface  may  be  accom- 
plished in  one  of  two  ways.  It  may  be  lowered  upon  a  truck 
and  removed  from  the  hoist-rope  and  taken  to  the  point  of  empty- 


HOISTING  MACHINERY,  ETC.  221 

ing,  meanwhile  being  replaced  by  an  empty  bucket  which  is 
lowered  below.  On  the  other  hand,  the  bucket  may  be  emptied 
at  the  surface  without  removal  from  the  rope.  In  the  becket 
and  high  under  the  bucket  the  shaftman  inserts  the  hook  of  a 
short  length  of  stationary  chain  The  tub  with  its  hoist-rope  is 


FIG.  87.— Loading-gate  for  Buckets  and  Cars. 

then  lowered  and  inverted,  its  contents  being  emptied  upon  a 
grating  at  the  side  of  the  shaft  or  into  a  car  which  is  run  over 
the  landing- doors  which  have  been  lowered  across  the  shaft.  In 
Fig.  89  is  another  device,  not  so  safe,  in  which  the  bale  is  pivoted 
a  little  below  the  centre  of  gravity  of  the  bucket  and  is  held  in 
position  by  a  loose  ring  on  the  bale,  slipping  over  the  pin  at  the 
upper  rim  of  the  bucket.  To  dump,  the  brakeman  merely  raises 
the  ring  and  allows  the  bucket  to  reverse.  It  is  easily  righted 
and  again  fastened. 

Aside  from  objection  to  hoisting  of  buckets  on  account  of  its 
low  speed  and  insecurity,  there  is  a  more  important  one  in  metal- 
mines,  namely,  the  production  of  considerable  powdered  mate- 
rial. This  soft  product,  though  it  may  reach  the  surface,  prob- 
ably never  reaches  the  smelter  and,  whatever  its  value,  is  lost. 


222 


MANUAL  OF  MINING. 


As  the  softer  minerals  are  also  the  richer,  one  can  readily  see 
the  necessity  for  avoiding  too  much  handling  of  the  mineral. 
In  addition  to  this  is  the  risk  of  collisions,  unless  each  tub  has 


FIGS.  88  and  89. — Mine  Buckets. 

its  own  hoist  compartment  with  smooth-lined  guides.  These 
slides  permit  of  more  rapid  hoisting,  but  the  damage  and 
accidents  arising  from  contact  with  buckets  when  the  end 
of  a  plank  protrudes  make  the  risk  and  cost  quite  consider- 
able. 

Shaft-guides. — To  facilitate  rapid  hoisting,  the  shafts  are 
fitted  with  conductors  or  guides,  along  which  the  conveyances 
move.  With  a  bucket-hoist  on  the  slopes  the  floor  is  planked 
with  4-inch  wooden  guides  spiked  on  either  side  of  a  runway, 
between  which  the  bucket  may  slide  in  its  passage.  The  ends 
are  finished  so  that  they  will  not  easily  loosen  and  cause 
accident.  .  Experiments  with  steel  rails  for  guides  have  proved 
them  far  too  noisy  for  comfort. 


HOISTING  MACHINERY,  ETC. 


223 


When  a  bucket  is  hoisted  in  a  shaft  it  swings  freely,  but  is 
guided  by  two  or  four  ropes  stretched  from  top  to  bottom;  these 
pass  through  holes  at  the  ends  of  two  or  four  arms  which  make  a 
horizontal  frame.  This  frame  rests  freely  above  the  shackles  of 


FIG.  90. 

the  hoist-rope  and  keeps  the  bucket  in  a  central  position  during 
all  times  of  the  hoist.  Cages  are  guided  by  two  4-inch  timber 
strips  which  are  bolted  to  the  buntons,  or  cross-pieces,  fixed 
across  the  pit.  These  are  securely  wedged  against  the  rock,  or 


£24 


MANUAL   OF  MINING. 


are  framed  into  the  timbering  of  the  shaft.    One  guide  on  each 
of  the  two  sides  is  sufficient  for  a  cage  compartment. 

Skips,  or  Gunboats,  are  commonly  seen  in  the  inclined  shafts 


''•--.-.  I    • 


FIG.  91. 


FIG.  92. 


for  hoisting  purposes.     They  consist  of  an  iron  box  weighing 
from  900  to  1500  Ibs.  set  on  four  wheels  held  by  bolts  riveted 


HOISTING  MACHINERY,  ETC.  22$ 

at  the  sides;  the  front  view  is  rectangular,  but  the  side  view  is 
trapezoidal  (Fig.  90).  The  inclined  end  is  uppermost  with  auto- 
matic dumpers,  while  in  those  discharging  by  door  it  is  below. 
The  hoist-rope  is  attached  to  the  bale,  which  rotates  on  a  pin 
passing  through  the  side  of  the  skip,  often  back  of  the  centre 
of  gravity,  so  as  to  dump  automatically.  The  charge  of  one 
or  two  car-loads  is  shovelled  or  chuted  into  the  skip,  which  emp. 
ties  at  the  surface  into  a  bin  or  on  grizzly.  In  one  variety  the 
contents  are  discharged  by  the  mouth  at  which  it  is  loaded,  while 
the  other  form  has  a  swinging  door  opening  at  the  upper  side. 
A  vertical  safety  skip  is  shown  in  Figs.  91  and  92. 

The  automatic  dump  is  simple;  the  rear  or  lower  wheels  are 
of  wider  face  than  the  fore  wheels.  As  the  dumping-plat  is 
reached,  the  guide-rails  on  which  the  wheels  have  been  travel- 
ling gradually  bend  to  horizontality,  and  these  the  front  wheels 
follow.  As  the  hoisting  continues,  the  wider  rear  wheels  catch 
and  roll  on  a  pair  of  outer  guides,  and  continue  up  the  slope. 
By  this  means  the  lower  end  is  elevated  and  the  skip  emptied. 

The  brake  is  generally  a  drag,  consisting  of  a  bar  about  4  feet 
long,  trailing  on  the  floor,  and  only  catching  if  the  skip  breaks 
away  on  its  up-trip.  Often  the  wheels  are  confined  between  two 
guides  on  each  side,  as,  for  instance,  where  the  direction  of  the 
slope  changes.  The  sole  objection  to  these  skips  is  the  double 
and  treble  handling  involved.  The  car  from  the  mill-hole  empties 
into  the  chute,  whence  the  skip  is  loaded,  and  at  the  surface 
the  reverse  operation  takes  place. 

The  Slope-carriage  dispenses  with  two  handlings,  taking  the 
car  at  once  to  the  surface.  This  is  simply  a  double  triangular 
frame,  large  enough  to  accommodate  a  car  with  two  rails  on 
its  horizontal  top,  and  two  wheels  on  each  hypothenuse.  A 
hook  or  lock  holds  the  car  while  riding.  For  convenience  the 
loading  and  unloading  gangways  are  not  on  the  same  level, 
the  track  for  "empties"  being  6  feet  higher  than  the  "full" 
track.  If  the  seam  is  too  thin  to  allow  of  this,  each  track  has 
a  curved  roadway  connection  with  the  gangway. 

The  head-room  required  for  the  slope  having  this  means  of 


226 


MANUAL  OF  MINING. 


haulage  implies  large  area  and  extensive  timbering,  particularly 
as  a  double  trackway  is  necessary  if  the  output  is  considerable. 
There  is  not  much  necessity  for  a  carriage,  except  in  slopes  over 
60°. 

The  Cage. — The  cage  is  essentially  a  wrought-iron  or  steel 
platform  for  the  cars  traversing  the  pit.    It  is  also  the  usual 


FIG.  93. — A  Safety  Cage  with  Keeps. 

means  of  transport  for  the  workmen  and  all  other  materials 
^between  the  surface  and  the  different  landing-stages  in  the  shaft. 
The  pit  timber,  workmen,  tools,  fodder  and  water  for  the  horses, 


HOISTING  MACHINERY   ETC.  227 

and  frequently  the  horses  themselves,  are  lowered  by  means  of  the 
cage.  Its  form  is  governed  by  the  shape  of  the  compartment  of  the 
shaft  in  which  it  runs,  and  its  size  by  the  number  of  cars.  Some 
cages  are  fitted  with  two  platforms  or  decks.  On  each  deck  are 
rails  for  one  or  two  cars,  and  so  placed  as  to  allow  of  those  on  the 
cage  being  displaced  by  others  to  be  hoisted  or  lowered.  The 
platform  is  held  on  the  sides  next  to  the  guides  by  two  stout  iron 
frames  united  at  the  top  by  a  cross-bar  to  which  the  hoist-rope 
is  attached  (Fig.  93) .  Two  iron  ears  at  the  top  and  bottom  on 
each  side  confine  the  cage  to  the  guides.  An  iron  roof  or  bonnet 
over  the  cross-bar  shields  the  passengers  from  falling  rocks. 

Some  cages  are  also  provided  with  extra  covers,  either  fixed 
or  adjustable,  above  the  tops  of  the  cages,  to  protect  the  men  while 
examining  or  repairing  the  shaft. 

An  automatic  device  controlling  the  delivery  of  the  cars  in 
and  out  of  the  cage  is  illustrated  in  n,  Fig.  84.  It  serves  for 
the  intermittent  delivery  of  cars  without  an  attendant. 

Securing  the  Cars  on  the  Cage. — There  are  various  modes  of 
keeping  the  car  secure  in  the  cage  during  its  ascent  or  descent. 
One  of  these  is  a  "fabe  bottom"  in  the  cage,  which  sinks  an 
inch  or  two  below  the  other  and  outer  portions  of  the  deck,  but 
is  lifted  when  the  cage  is  at  a  landing-station.  The  deck  floor 
is  then  on  one  level,  allowing  the  tub  to  be  changed. 

Another  device  is  a  bar  running  across  the  cage,  and  at  either 
end  is  placed  a  short  lever,  which  turns  down  or  up  on  being 
pushed;  when  down  it  covers  the  ends  of  the  tubs  and  prevents 
their  moving;  when  up  it  allows  them  to  be  changed.  Some- 
times a  latch-lock  on  the  floor  secures  the  cars  in  place. 

Multiple-deck  Cages.— Single-deck  cages  are  almost  exclu- 
sively used  in  America,  and  are  sufficient  except  for  narrow,  deep 
shafts,  when  heavier  loads  are  necessary  for  large  output.  Then 
cages  with  two  or  more  platforms  are  used.  The  landing-stages 
are  arranged  in  the  same  number  of  tiers,  from  and  upon  which 
cars  are  simultaneously  run,  without  moving  the  cage.  With 
ample  facilities  for  "decking"  the  cars,  the  saving  in  the  trip- 
time,  per  car,  increases  the  capacity  of  the  plant.  The  necessarily 


228 


MANUAL  OF  MINING. 


complicated  underground  stations  are  an  objection  to  multiple- 
deck  cages.  The  double-platform 
carcage  is  relatively  lighter  than 
the  single- deck  two-car  cage. 

For  inclines  of  uniform  slope, 
as  well  as  for  vertical  shafts,  cages 
may  be  had,  the  platform  being 
hung  on  an  adjustable  lever;  but 
the  slope  carriage  would  be  better. 
Safety  Clutches. — Various  con- 
trivances are  proposed  for  guarding 
against  accidents  resulting  from  the 
breaking  of  the  rope.  The  sudden 
starting  and  stopping  of  the  hoist 
produces  a  shock  which  strains  the 
cable  more  severely  than  does  the 
mere  weight.  Without  care  in  hoist- 
ing and  rigid  timbering  the  rope 
cannot  be  insured  against  the  jerks. 
To  avoid  the  accident  sure  to  follow, 
some  form  of  safety  appliance  must 
be  thrown  into  action,  to  arrest  the 
fall  of  the  cage  or  skip.  The  safety 
catches  differ  in  design  and  effi- 
ciency, but  depend  generally  upon 
a  spring,  so  held  between  the  rope 
and  the  cage  as  to  be  compressed, 
while  the  weight  of  the  cage  strains 
the  rope,  but  acts  on  a  clutch  that 
grasps  the  guides  and  stops  the 
fall,  if  relaxed  by  rupture  of  the 
rope  or  otherwise.  The  clutch  is 
either  a  pair  of  sharp-pointed  steel 
levers,  which  are  thrust  outward 
into  the  timbers,  or  a  serrated 
cam,  the  wider  part  of  which  will 


FIG.  94. — A  Double-deck  Cage. 


HOISTING  MACHINERY,   ETC.  229 

be  turned  against  the  guides  and  clutch  them  on  either  side 
(Fig.  93).  The  heavier  the  weight  the  stronger  will  be  the 
grip  after  they  have  once  taken  hold. 

In  Callow's  attachment,  13,  Fig.  84,  a  heavy  weighted  lever 
on  top  of  the  cage,  held  up  by  a  spring  only  so  long  as  the  rope  is 
intact,  is  set  into  action  and  engages  the  clutches  upon  the  guides 
when  the  rope  breaks.  Many  a  life  has  been  saved  by  them, 
but  in  many  instances  also  they  have  failed  to  operate.  A  momen- 
tary check,  or  any  sudden  change  in  the  speed,  often  unnecessarily 
throws  them  into  action.  On  the  other  hand,  they  are  rarely, 
if  ever,  in  order  when  the  emergency  arises,  or  the  guides  are  so 
wet  and  dirty  that  the  clutches  fail  to  catch,  if  the  momentum  of 
the  falling  cage  is  great;  besides,  they  are  costly  and  troublesome. 
Though  useful  adjuncts,  which  the  law  requires,  yet  it  is  not  sur- 
prising that  the  distrust  of  them  is  strong;  they  give  rise  to  other 
and  more  serious  causes  of  alarm.  The  tendency  following  their  use 
is  to  rely  upon  them  solely  and  to  neglect  the  rope.  Others  deem 
them  of  no  value  and  treat  them  accordingly.  Fortunately,  the 
rope  more  often  breaks  at  the  moment  of  starting  at  the  bottom 
than  at  any  other  time,  and  the  point  of  rupture  is  where  the 
rope  enters  the  socket.  If,  however,  the  rope  snaps  as  it  turns  the 
sheave — and  this  is  of  common  occurrence — there  is  nothing  to 
prevent  the  inevitable  and  frightful  calamity  that .  follows — the 
entire  length  of  the  rope  falls  and  crushes  cage  and  cbntents.  A 
simple  appliance  might  be  introduced  at  the  top  to  grasp  the  rope. 

Landing-doors  at  Shaft  Stations. — The  landing  of  the  buckets 
is  effected  on  a  hinged  door,  of  double  2-inch  planks,  lined  on  top 
with  iron.  This  is  swung  against  the  far  side  of  the  compartment, 
closes  it  completely,  and  standing  as  it  does  at  45°,  the  bucket 
slides  into  the  drift.  When  not  in  use  it  is  hooked  upright,  closing 
the  mouth  of  the  drift  and  leaving  the  hoistway  clear.  That 
miners  may  go  from  one  drift  to  the  opposite,  an  escapement 
roadway  is  provided  around  the  shaft  in  the  rock.  At  the  mouth 
of  the  shaft  are  two  heavy  doors,  opened  for  or  by  the  rising 
bucket  and  closed  at  all  other  times.  These  are  convenient, 
and  prevent  accidents  arising  from  stones  dropping  down  the 


230  MANUAL  OF  MINING. 

shaft,  especially  while  dumping  the  bucket.  A  similar  arrangement 
in  slopes — a  movable  drawbridge,  lowered  at  pleasure — receives 
the  car  for  attachment  to,  or  after  detachment  from,  the  rope,  and 
closes  the  gangway  when  the  car  is  above  the  plat. 

Cage-chairs. — At  the  surface  and  at  every  landing- station  the 
cage  is  supported  upon  chairs,  known  also  as  "keeps,"  "fans,"  or 
"shuts."  They  are  counterbalanced  levers  (Fig.  93),  which  pro- 
ject into  the  shaft,  giving  way  for  the  ascent  of  the  cage,  and 
immediately  resume  their  position  after  the  cage  passes  them. 
The  latter  rests  upon  them  during  loading.  To  lower  the  cage 
an  attendant  raises  them  out  of  the  way  while  the  cage  is  being 
lifted  off. 

With  double- decked  cages  the  top  deck  of  the  cage  at  the 
shaft  bottom  rests  on  keeps,  while  the  bottom  of  the  other  cage 
stands  on  similar  supports.  When  the  cars  have  been  changed 
a  signal  is  given,  the  cages  are  lifted,  while  the  attendant  at  the 
surface  pulls  them  clear  of  the  shaft  until  the  bottom  deck  has 
been  lowered  below  them.  Releasing  the  handle  thrusts  them 
into  the  shaft  to  hold  the  cage  at  its  top  deck. 

Those  at  the  shaft  bottom  are  necessarily  handled  differently. 
As  the  loaded  cage  leaves  the  shaft  bottom  the  attendant  there 
pulls  the  handle  of  the  keeps  back  and  secures  it,  thus  preventing 
them  from  protruding  into  the  pit.  Holding  the  lever  working 
the  keeps,  when  the  lower  deck  of  the  cage  has  passed  below 
their  level,  he  allows  them  to  spring  out  to  the  support  of  the  top 
deck.  The  car  is  then  changed,  while  the  lower  deck  of  the 
other  cage  is  being  changed  at  the  surface.  The  keep-handle 
will  not  require  further  attention  from  the  attendant  below  until 
the  cage  has  left  for  the  surface,  when  he  secures  the  handle  back 
in  its  place. 

Stauss's  patent  keeps,  shown  in  15,  Fig.  84,  have  been  de- 
signed to  lower  the  cage  into  the  pit  without  a  previous  lift.  They 
are  fixed  on  wooden  spring  cantilevers,  to  take  the  shock  of  the 
cage  when  it  is  lowered  on  them.  At  16,  Fig.  84,  is  the  end  view 
of  another  style  of  lowering-keep.  These  reduce  the  wear  and 
tear  of  the  winding  ropes  and  engines.  The  cage  can  be  low- 


HOISTING  MACHINERY,  ETC. 


23I 


ered  immediately  without  the  lifting  jerks,  which  are  so  grea  at 
cause  of  deterioration  of  ropes  and  engines.  When  these  keeps 
are  used  the  winding  rope  must  of  course  be  carefully  adjusted 
in  length  so  that  the  cage  does  not  fall  after  drawing  back  the 
keeps.  The  sinking  of  the  cage  should  not  amount  to  more 
than  the  slack  of  the  rope  when  unloaded.  The  length  of  the 
rope  can  easily  be  adjusted  by  means  of  cage-adjusting  hangers, 
which  are  essentially  of  the  turnbuckle  type. 

Self-dumping  Cages. — The  self-dumping  cage  (Fig.  95)  is  a 
bucket  supported  below  the  cage,  whose  floor  is  divided  and 
hung  at  the  sides  like  two  doors.  Below  it  is  loaded  from  cars, 
and  at  the  surface  it  is  dumped  very  easily.  Unless  the  coal  is 


FIG.  95. — A  Self-dumping  Cage. 

to  be  coked,  tender  coal  cannot  be  treated  in  this  way.  Where 
all  the  coal  is  to  be  coked,  or  at  a  small  mine  marketing  its  coal, 
self-dumping  cages  are  used,  only  one  man  being  required  for 
the  work.  The  capacity  of  a  pair  of  them  is  limited  to  1000  tons 
of  run-of-mine  coal  if  screened  over  one  set  of  bars. 

Mine-cars.  —  The  dimensions  of  cars  are  reduced  to  a  mini- 
mum commensurate  with  the  size  of  the  haulage-way,  and  they 


*32  MANUAL   OF  MINING. 

must  be  large  for  purposes  of  economy.  To  obtain  the  highest 
efficiency  they  are  made  as  large  as  the  natural  conditions  of 
grade  in  the  haulage-ways  will  allow.  Their  weight  is  reduced 
to  a  minimum  in  order  to  economize  on  haulage-power.  Indeed, 
the  amount  of  power  available  determines  their  maximum  allow- 
able gross  weight  and  grades  suitable  for  the  haulage-ways. 

In  small  mines  where  man-power  is  employed  for  tramming, 
the  limit  of  weight  and  size  is  fixed  by  the  inclination  of  the  return 
track;  so,  too,  in  buggy-rooms  the  weight  and  capacity  of  the 
car  depends  upon  the  power  there  employed.  These  cars  are 


FIG.  96. — A  Metal  Mine-car. 

therefore,  limited  to  small  dimensions.  Coal-cars  are  much 
larger  and  heavier  when  mechanical  systems  of  haulage  can  be 
utilized.  If  the  conditions  permit  automatic  delivery  of  the  cars 
from  the  face  to  the  locomotive  or  the  haulage-way,  their  weight 
is  that  which  the  mule  or  horse  can  conveniently  return  empty 
to  the  face.  Evidently  the  car  should  be  as  light  as  is  com- 
patible with  its  strength.  Its  weight  is  about  one  half  that  of  its 
contents.  In  metal-mines  the  proportion  is  somewhat  greater. 
The  car  weighs  from  600  Ibs.  to  1500  Ibs.  Some  coal-cars  carry 
as  much  as  120  cu.  ft.  of  mineral,  but  the  capacity  of  those  in 
metal-mines  is  less  than  30  cu.  ft.  For  strength,  compactness, 


HOISTING  MACHINERY,  ETC. 
T" 

~T 


233 


234  MANUAL   OF  MINING. 

and  tightness  sheet-steel  car-bodies  are  used  where  the  mineral  is 
hard,  as  in  metalliferous  mines.  Otherwise  a  wooden  box  is  em- 
ployed. If  the  mineral  is  heavy  but  soft,  it  is  necessary  to  employ 
a  sheet-steel  boxing  in  preference  to  wood 

The  width  of  the  cars  depends  upon  the  gauge  and  its  set. 
The  bodies  resting  on  the  axles  are  preferred,  but  may  not  be 
advantageous  because  of  the  wide  gauge  of  track.  On  the  other 
hand,  with  a  narrow-gauge  track  and  given  capacity  the  car- 
body  is  elevated  and  its  centre  of  gravity  too  high.  A  compro- 
mise is  frequently  taken  by  which  a  low  wide  car  of  great  capac- 
ity may  be  raised  on  a  narrow  track.  In  this  case  the  axles  are 
elbowed  for  large  wheels  and  the  car-body  is  set  down  on  them, 
but  is  widened  above  the  wheels  to  full  width  of  the  track.  Their 
width  must  provide  ample  clearance  for  passing  men  in  the  haul- 
age, and  their  length  allow  for  space  in  the  cage  compartment. 
The  length  is  limited  by  sharpness  of  the  curves.  In  the  collieries 
the  maximum  is  9  feet,  and  in  metal-mines  it  is  not  over  6  feet. 

Their  height  depends  upon  the  conditions  of  loading.  In 
metalliferous  mines  the  cars  run  only  in  main  haulways,  and  are 
filled  from  chutes,  provided  with  a  spout  and  gate,  easily  manip- 
ulated at  the  bottom;  if  also  hoisted  on  a  cage,  their  height  is  a 
matter  of  indifference.  When  the  seam  is  thick  and  flat  and  the 
roof  good,  they  are  carried  to  the  face  of  the  work,  in  which  case 
they  are  filled  by  hand.  If  so,  or  if  raised  on  carriage,  the  diffi- 
culties and  expense  are  greater  with  a  high  car  To  shovel  one 
ton  into  a  3-foot  car  requires  over  7300  ft.-lbs. ;  into  a  car  4  feet 
high,  9500.  The  average  man  can  exert  a  continuous  shovelling 
effect  of  28,100  ft.-lbs.  per  hour.  Allowing  for  the  weight  of  the 
shovel,  delays,  throwing  the  mineral  forward,  a  shoveller  may 
load  about  20  and  14  tons,  respectively,  in  the  cars  per  shift. 
Even  for  a  medium  output  the  economy  is  manifest.  In  metal- 
liferous mines  a  low  car  is  used,  but  in  collieries  heights  of  4  feet 
9  inches  and  over  are  common.  For  stability,  too,  a  low  car  is 
desirable. 

Car  Details.— The  steel  cars  are  purchased  of  the  manu- 
facturers and  may  be  had  to  order  of  any  size  and  design  for  the 


HOISTING  MACHINERY,  ETC. 


235 




II 

1     l!    ;]         -- 

•"•Ljj-jjJ  <®  \       1 

I 


r 


MANUAL   OF  MINING. 


111 


HOISTING  MACHINERY,  ETC.  237 

given  requirements.  They  are  of  steel  plate  \  inch  thick,  with 
their  sides  and  back  end  in  one  piece.  The  bottom  plate  is 
fastened  to  the  sides  by  iron  angle-arms.  The  standard  coal- 
cars  are  of  the  shape,  size,  and  details  as  shown  in  Figs.  96  to  100. 

The  metal  cars  are  supported  on  one  four-wheeled  truck. 
The  wooden  cars,  on  two  trucks,  are  almost  all  of  them  arranged 
with  a  swing-door  at  the  end  and  provided  with  a  trip-catch  for 
a  quick  release.  It  is  rare  for  companies  to  make  their  own 
cars,  though  all  make  their  own  repairs. 

The  Wheels  are  as  large  as  circumstances  will  permit  (the 
larger  the  wheels  and  the  smaller  the  axles,  the  less  is  the  fric- 
tion). The  wheels  may  revolve  loosely  on  the  round  or  the 
square  axle,  or  they  may  be  fixed  to  the  axle  and  revolve  with 
it.  Some  are  capped  with  a  recess  in  the  hub,  to  receive  the 
collar  on  the  axle,  and  thus  prevent  admission  of  grit  (Fig.  100). 
They  may  be  "inside"  (below)  or  "outside"  (beyond)  the 
body  of  the  car.  As  to  the  relative  merits  of  the  inside  and  out- 
side, or  loose,  wheels,  it  must  be  admitted  that  engineers  are 
not  unanimous,  though  the  former  have  the  larger  number  of  ad- 
herents. Outside  wheels  are  more  easily  oiled,  are  cheaper,  and 
admit  of  the  body  of  the  car  being  set  lower  down;  they  do  not 
run  as  smooth  or  last  as  long  as  those  fixed  under  the  body  of 
the  car.  Loose-wheel  cars  may  be  better  for  short  roads  with 
sharp  curves,  but  they  are  harder  to  pull.  With  fixed  wheels,  one- 
of  the  mutually  dependent  wheels,  in  travelling  about  curves,  must 
slide.  For  this  reason,  and  because  they  are  easily  lubricated,  loose- 
wheels,  or  cone-fixed  wheels,  are  preferred  by  many.  A  great 
many  mines  have  abandoned  loose  wheels  after  careful  trial. 
At  the  Drifton  anthracite  mine  a  compromise  is  effected  by- 
using  a  pair  of  fixed  and  a  pair  of  loose  wheels. 

The  coal-car  wheels  are  of  cast  iron,  between  16  and  18". 
inches  diameter,  and  those  of  ore-cars  about  one  half  that  and' 
solid.  The  former  have  hub  and  arms  to  allow  of  "spragging." 

The  wheel-base  of  the  car-trucks  is  a  trifle  greater  in  lengths 
than  the  width  of  the  track.  Its  width  is  f  inch  less  than,  the- 
track  gauge. 


338  MANUAL  OF  MINING 

The  life  of  a  car  depends  upon  conditions  too  numerous  to 
detail  here,  but  it  may  be  given  as  averaging  three  years  in  coal- 
mines and  a  trifle  longer  in  metal-mines.  They  all  grow  larger 
'with  age.  The  wooden  cars  become  loose  and  swell;  the  iron 
ones  are  quickly  battered  out  of  shape. 

Brakes  are  provided  on  all  coal -cars  which  go  to  the  face  of 
the  workings.  Otherwise  they  are  not  required. 

Car  Resistance. — The  frictional  resistances  offered  by  the 
wheels  at  their  circumference  on  the  track  and  at  their  journals 
amount  to  about  2  per  cent  of  their  total  weight.  This  quantity 
may  be  smaller  where  care  has  been  given  to  their  construction, 
as,  for  example,  those  which  are  employed  in  connection  with  the 
electric  mine-haulage.  This  friction  can  be  reduced  by  the 
liberal  use  of  lubricants  and  care  in  the  finishing  of  the  journals. 
Jf  the  latter  are  boxed  and  kept  free  from  dust,  the  frictional 
resistance  will  be  materially  reduced. 

The  lubrication  of  the  bearings  may  be  obtained  in  the  ordinary 
-way  by  the  use  of  oil -soaked  waste  in  the  boxes  forced  against 
the  journals.  Self-oiling  journals  would  further  reduce  the  fric- 
tion. They  are  inexpensive,  and  are  desirable  if  saving  in  power 
is  to  be  considered.  By  their  use  one  third  of  a  pint  of  oil  will 
last  from  two  to  four  months,  according  to  its  quality  and  the  dis- 
lance  traversed.  The  saving  in  power  is  $10  per  year  per  car, 
compared  with  the  inefficient  methods.  This  includes  also  the  cost 
of  axle  replacements.  Self-oilers  are  also  economical  of  lubri- 
cant, as  compared  with  the  crude  method  ordinarily  employed 
of  pouring  oil  into  a  box  and  bearing.  The  oil  used  is  usually 
one  of  the  heavy  varieties  of  lubricant  with  the  addition  of  some 
axle-grease. 

Lubricants. — An  efficient  lubricator  must  have  sufficient 
body  to  keep  the  surfaces  between  which  it  is  interposed  from 
coming  into  contact.  It  must  have  the  greatest  fluidity  con- 
sistent with  its  needed  body.  A  small  coefficient  of  friction  is 
necessary,  as  well  as  a  maximum  capacity  for  conducting  the 
heat.  It  should  be  free  from  any  tendency  to  oxidize  or  gum, 
and  free  from  acid- 


HOISTING  MACHINERY,  ETC. 


239 


240 


MANUAL  OF  MINING. 


Testing  Lubricating-oils. — The  engineer  should  have  means 
for  testing,  as  oils  intended  for  lubricating  purposes  are  often 
adulterated  with  foreign  and  even  injurious  matter.  A  hydrom- 
eter and  flash-test  cup  are  sufficient  for  this  purpose. 

THE  FLASH-POINT  AND  SPECIFIC  GRAVITY,  BEAUME,  OF  LUBRICATING  OILS. 


>, 

1 

A 

£ 

.,- 

. 

Is 

P 

I1 

P 

II 

a 

11 

0 

Ac* 

ill 

Castor-oil  (pure). 
Linseed-oil  

0.9639 
.9299 

15 
19 

5°5 

Tallow-oil  
Heavy  engine-oil  .  . 

0.9080 
.9032 

24-5 
25-5 

540 

414 

Cottonseed-oil.  .  . 

.9210 

jg 

518 

Castor  machine-oil 

.8919 

27 

324 

Lard-oil  

•9*75 

23 

505 

Light  engine-oil.  .  . 

.8917 

27 

342 

Rapeseed-oil  

•9155 

23 

Sperni-oil  

.8815 

29 

Cvlinder-oil 

Spindle-oil  

.8^88 

3? 

?  j  2 

° 

The  Hydrometer. —All  oils  are  lighter  than  water  and  their 
purity  may  be  ascertained  by  the  use  of  a  Beaume  hydrometer. 
If  a  contract  calls  for  a  heavy  engine-oil  of  25°  Beaume,  the  hy- 
drometer at  once  reveals  whether  the  same  is  received.  If  adul- 
terated, a  lower  Beaume  reading  reveals  the  blending.  The 
adulterants  used  are,  of  course,  cheaper  than  the  pure  oil.  A 
"lard-oil,"  for  example,  of  23°  Beaumf,  sp.  gr.  27,  must  have 
been  adulterated  with  a  lighter  oil,  and  the  only  cheaper  ones 
possible  are  the  light  petroleum  or  the  neutral  oils. 

Viscosity  Test. — The  viscosity  and  gumming  tendency  of  oils 
may  be  detected  by  noting  the  time  required  to  traverse  a  given 
distance  down  an  inclined  plane,  and  the  time  when  this  ceases. 
The  oil  remaining  limpid  the  longest  has  the  lowest  gumming 
tendency,  and  that  which  flows  the  farthest  is  the  most  viscous. 
A  nine-days'  test  of  oils  gave  the  following  results:  A  heavy 
mineral-oil,  for  engine  use,  traversed  87  inches  on  the  ninth 
day;  common  sperm-oil,  63  inches  on  the  ninth  day;  olive-oil, 
22  inches  on  the  ninth  day;  linseed-oil,  18  inches  on  the  seventh 
day;  and  lard-oil,  12  inches  on  the  fifth  day.  In  each  case  the 
day  given  is  that  on  which  the  oil  ceased  to  travel. 

The  simplest  method  of  testing  the  fluidity  is  to  dip  blotting- 
paper  into  the  oil  and  hold  it  up  to  drain.  An  oil  of  good  fluidity 


HOISTING  MACHINERY,  ETC.  241 

presents  symmetrical  drops.  One  of  high  viscosity  spreads 
readily  over  the  paper.  The  rate  of  gumming  may  also  be 
ascertained  by  its  retention  on  the  paper,  while  at  high  tem- 
perature, for  some  hours. 

A  Ring  Test. — An  oil  which  is  composite  in  nature  will  show 
on  a  clean  sheet  of  blotting-paper  two  or  more  rings,  the  outer 
one  indicating  the  poorest  oil.  The  central  denned  ring,  dark 
in  color,  is  due  to  the  heavier  oil.  If  the  rings  disappear,  it 
is  mineral-oil.  Paraffine  stock  in  the  oil  will  be  revealed  by  a 
well-discerned  ridge  near  the  centre.  A  light  hydrocarbon  oil 
does  not  give  any  permanent  translucency  to  the  paper. 

The  Flash-point  Test. — This  consists  in  finding  the  tempera- 
ture which  an  oil  will  endure  without  being  vaporized  and  ignited. 
An  oil  for  cylinders  should  have  a  high  flash-point;  those  not 
intended  for  lubricating  steam  surfaces  may  have  a  much  lower 
flash-point.  The  flash-point  can  be  determined  by  heating 
a  sample  of  the  oil  in  a  cup  over  an  alcohol-lamp,  on  a  tripod. 
A  thermometer  inserted  in  the  oil  will  note  the  degree  at  which 
ignition  takes  place.  Care  should  be  taken  that  the  thermometer 
touch  neither  the  bottom  nor  the  sides  of  the  cup.  As  the  oil 
reaches  300°  F.  a  match  or  lighted  twine  should  be  passed  across 
the  surface  of  the  oil  until  a  blue  flame  shows  the  time  of  ignition. 
Lard-oil  has  a  high  flash-test.  If  the  sample  submitted  burns 
below  500°,  it  is  adulterated  with  cottonseed  or  petroleum;  lin- 
seed-oil will  have  its  flash-test  reduced  by  the  same  adulterants. 

Fire-proof  Oils. — These  are  valuable  on  bearings  inclined  to 
heat.  They  derive  their  virtue  from  the  fact  that  they  contain 
considerable  water,  held  in  suspension  in  the  oil  by  means  of  a 
mineral  soap.  When  applied  to  heated  bearings  the  water 
serves  to  absorb  the  frictional  heat,  leaving  the  oil  for  lubrication. 

Light-running  machinery  is  lubricated  with  oil  of  300°  flash- 
test,  heavy  machinery  with  oil  of  400°,  and  engines  with  oil  of 

550°. 

Oil  intended  for  the  bearings  of  motors  and  generators  ex- 
posed to  low  temperature  should  be  limpid  and  free-flowing. 

Cylinder  lubrication  is  best  obtained  from  a  petroleum-oil  of 


242  MANUAL   OF  MINING. 

high  flash-test  (530°)  and  viscosity.  It  is  compounded  with 
5  per  cent  of  tallow-oil  if  the  steam  is  wet. 

The  air-cylinders  of  air-compressors  and  of  gas-engines  should 
be  lubricated  only  by  a  pure  petroleum-oil. 

Geared  wheels  are  oiled  by  some  preparation  composed  of  tar, 
wax,  and  similar  substances  combined  with  rosin,  which,  when 
applied  to  the  teeth  in  a  molten  state,  will  cool  to  a  tough,  elastic 
coating,  reducing  friction  and  noise. 

Handling  Cars  at  the  Tipple.  —  To  facilitate  the  delivery 
of  the  mineral  at  a  minimum  expense,  and  with  as  little  hand- 
ling as  possible,  the  landing-station  or  the  tipple  is  fitted  with 
quick-dumping  appliances  for  the  cars  and  suitable  grades  for 
their  prompt  return.  The  level  of  its  floor  is  below  that  of  the 
mine  mouth  or  the  landing-station  in  the  derrick,  and  far 
enough  above  the  railroad  track  or  the  point  of  shipment  to  per- 
mit the  mineral  to  have  a  continual  descent  between  the  time  of 
dumping  and  the  time  of  shipment.  The  tracks  to  the  tipple  are 
laid  favorably  for  the  load,  and  the  character  of  the  conducting 
line  between  them  will  depend  upon  the  distance  and  the  slope 
of  the  ground.  In  the  case  of  a  slope  mine  the  loaded  cars  or 
the  train  can  be  hauled  directly  to  the  tipple  and  returned  thence 
by  gravity.  Any  grade  exceeding  one  per  cent  will  allow  the 
empty  car  to  return. 

Dumping  Cars. — The  dumping  of  cars  may  be  accomplished 
by  opening  one  side  or  the  end  of  the  car  by  inclining  the  car 
on  its  axle  or  by  tipping  it  upon  some  form  of  cradle.  Mineral 
which  is  to  be  subsequently  broken  or  sorted  is  dumped  upon 
a  screen  or  the  grating  to  the  sorting-floor.  Some  tipples  are 
provided  also  with  self-indicating  weighing-machines,  by  which, 
as  the  car  is  emptied,  the  machine  registers  its  weight,  turns 
over,  shoots  out  the  coal,  and  returns  to  its  normal  position  for 
another  load. 

The  iron-body  car  of  metal-mines  is  set  on  a  swivel  near  the 
centre  and  held  in  position  by  a  lever.  When  it  is  to  be  emptied 
the  lever  is  unhooked,  the  car  turned  to  position,  the  door  raised, 
and  the  car  tilts  because  of  the  slight  overhang  at  the  discharging 


HOISTING  MACHINERY,  ETC. 


243 


end.     It  empties  without  further  effort  from  the  trammer  (Fig. 
101).    It  is  fastened  by  the  latch  as  soon  as  the  car  body  has  been 


FIG.  loi.— A  Steel  Ore-car. 


swung   into   position   on   its   truck.     Cars   intended   for  tunnel 
use  empty  at  the  end,  the  door  being  hung  by  strap  hinges  from 


FIG.  102. — A  Side  Dump-car 

an  iron  rod  at  the  top.     It  latches  in  the  same  way  as  mentioned. 
Another  variety  of  tunnel   car  consists  of  double   iron  frame, 


.?44 


MANUAL  OF  MINING. 


pivoted  at  the  middle  of  the  top  line  on  the  side.  It  is  on  two 
trucks  for  dumping.  The  latch,  located  at  each  side,  is  raised, 
the  car  opens  at  the  middle  and  empties  between  the  trucks 
(Fig.  103).  The  doors  on  the  ends  of  coal-mine  cars  are  locked 
by  latches  similar  to  those  of  the  iron-frame  cars.  In  Figs. 


Cost  $100  + 


Wt.  900-1000  IDS. 
Capacity  %  cord  -  2  Tons 

FIG.  103. — Tunnel  Dump-car. 

105  and  1 06  are  illustrated  some  typical  latches  used  in  the  coal 
regions. 

The  empty  cars  can  be  returned  to  the  mine  by  being  raised 
by  a  short  chain-haulage  to  an  elevation  sufficient  for  them  ta 
return  by  gravity  to  the  mouth  of  the  mine.  A  chain-haul  can 
be  used  to  return  the  empty  cars  to  a  "kick-back"  beyond  the 
shaft  if  the  distance  from  the  dump  will  permit.  This  is  then 
worked  like  the  Ramsay  caging  apparatus,  but  is  much  cheaper 
than  the  latter. 

Cross-overs. — Various  combinations  of  tracks  and  machinery 
are  employed  to  expedite  the  loading  and  unloading  of  cars  on 
the  cage  and  economize  time  at  landing-stations.  The  plan  is  to 
displace  the  empty  car  upon  the  cage  by  one  to  be  placed  thereon. 
For  example,  the  loaded  car  pushed  by  hand  upon  a  cage  bumps 


HOISTING  MACHINERY,  ETC. 


245 


off  the  empty  car  at  the  bottom  of  the  shaft,  or  the  empcy  car  to 
be  lowered  into  the  mine  is  made  to  replace  the  loaded  car,  which 
is  simultaneously  run  off.  The  automatic  cross-overs  are  of  sev- 
eral types  and  very  effective.  The  usual  plan  consists  in  having 
the  car  approaching  the  dump  release  dogs  holding  the  empty 
car,  and  allow  it  to  pass  over  the  dump  to  the  "kick-back"  (Fig. 
104),  whence  it  may  be  returned  by  the  track  on  the  left. 

The  Ramsay  Caging  Apparatus. — This  consists  essentially  of 
two  steam-rams  placed  back  of  the  shaft  and  two  transfer-tracks, 
running  on  a  track  across  the  tipple  in  the  rear  of  the  head-frame, 
which  are  operated  by  a  steam-cylinder.  Its  operation  is  as 
follows : 


FIG.  104. — The  Ramsay  Caging  Apparatus. 

After  a  car  has  been  dumped  it  runs  back  by  gravity  past  one 
side  of  the  head-frame  into  one  of  the  transfer-tracks,  which  is 
then  moved  by  means  of  the  steam-cylinder  to  the  rear  of  the  com- 
partment where  the  next  loaded  car  is  coming  up.  When  the 
cage  with  the  loaded  car  is  at  the  landing,  the  empty  car  is  pushed 
by  means  of  one  of  the  steam-rams  against  the  loaded  car,  which 
is  then  taken  off  the  cage  and  the  empty  car  left  on  the  cage 
ready  to  descend,  so  that  the  loaded  car  is  shoved  from  the  cage 
and  the  empty  one  pushed  on  at  one  operation.  When  one 
transfer-track  is  back  of  the  frame  with  its  empty  car,  the  other 
track  is  in  position  to  receive  the  next  empty  car  as  it  comes 
from  the  dump. 


246 


MANUAL  OF  MINING. 


HOISTING  MACHINERY,  ETC. 


o 
ft 

3 

lo,    a 


CO 


u 


248 


MANUAL   OF  MINING. 


Cradle-dumps.  —  Various  forms  of  cradles  are  provided  at 
dumping-stations  to  facilitate  the  emptying  of  cars  with  or  with- 
out hinged  doors.  These  consist  of  some  form  of  balance-frame 
upon  which  the  loaded  car  is  run  and  held.  When  the  catch  is 
released  the  frame,  with  its  car,  tilts  and  empties  the  contents  as 
soon  as  the  centre  of  gravity  has  passed  slightly  beyond  the 
pivot,  or  axle,  of  support.  Another  form  (Fig.  107),  which  is  the 


FIG.  107. 

Behr  device,  is  much  used.  The  car  is  held  on  a  platform  and 
revolves  with  it  around  the  point  of  support,  as  indicated  by  the 
dotted  lines.  The  car  in  this  device  is  almost  completely  in- 
verted and  thus  emptied.  Another  form  of  cradle,  used  with 


HOISTING  MACHINERY,  ETC.  249 

larger  cars  having  no  swinging  doors,  consists  of  a  skeleton  barrel 
having  tracks  in  such  a  position  as  to  receive  the  car,  which  is 
then  completely  enclosed.  When  the  trap  is  released,  the  barrel  is 
revolved  by  gear-wheels,  and  the  car  is  inverted.  In  the  anthra- 
cite mines  are  employed  different  methods  of  dump,  as  illustrated 
in  Figs.  1 08  and  109.  The  shock  of  the  loaded  car  running 
upon  the  truck  carriage  supported  on  the  gravity  tracks  causes 
the  front  wheels  of  the  latter  to  roll  down  the  curve,  while  the 
end  wheels  are  elevated  and  the  car  is  tipped  sufficiently  to  empty 
its  contents,  the  door  having  been  previously  released. 

Means  of  Access  for  the  Men.— For  purposes  of  ingress  and 
egress  mines  are  provided  with  ladders  or  man-engines,  where 
the  cage  or  bucket  is  not  used.  The  statutes  of  the  several  States 
contain  various  requirements  for  the  accommodation  of  the  men. 
Some  require  the  maintenance  of  substantial  ladders  in  a  separate 
compartment  as  the  sole  means  to  be  used  by  the  men  for  entry 
and  exit.  In  other  States  operators  are  relieved  of  the  necessity 
of  keeping  up  a  ladder-way,  if  safety- carriages  are  employed. 
The  laws  of  many  States  forbid  the  use  of  buckets  by  the  miners, 
while  the  general  tendency  in  all  regions  is  to  insist  upon  two 
well-equipped  escapement-ways. 

If  the  angle  of  entry  is  below  30°,  no  special  provision  is 
necessary.  The  mudsills  of  the  timbering  break  the  descent 
into  sufficiently  convenient  steps.  For  steeper  angles,  up  to 
about  60°,  some  variety  of  tread  is  necessary.  When  the  pitch 
exceeds  this  the  compartment  must  be  provided  with  ladders 
isolated  from  the  hoistway.  They  should  be  inclined,  uni- 
'form  in  direction,  at  an  angle  of  not  less  than  10°  from  the 
vertical,  to  diminish  the  fatigue  of  climbing  and  enable  the 
men  to  carry  tools  with  them.  At  equal  distances  down  the 
ladder-way  (20  to  40  feet  down  a  vertical  shaft,  and  at  greater 
distances  on  an  incline)  platforms  are  built  of  2  X6  beams  and 
2-inch  planks,  closing  it,  except  for  a  manhole,  at  the  foot-wall 
end.  The  ladders  extend  up  through  the  manhole,  and  are 
fastened  by  staples  or  toe-nailed  to  the  shaft-timbers,  and  rest 
on  the  far  side  of  the  plats.  They  are  made  of  2  X6  standards, 


250 


MANUAL  OF  MINING. 


*«* 


7\£ 


• 

« 


FIG.  108.— Anthracite  Car-dump. 


HOISTING  MACHINERY,  ETC. 


1 8  inches  apart,  with  iron  or  wooden  rounds  or  rectangular  slats 
12  inches  apart.     The  last-named  are  cheaper,  last  longer,  and 


give  better    toe-hold  than  wooden  rounds,  which,  in  turn,  are 
easier  to  use  than  the  more  durable  iron. 

In  slopes    having  an  inclination  of  less  than  60°  steps  and 
rails  are  provided,  the  tread  being  6  inches  and  the  slope  dis- 


252  MANUAL   OF  MINING. 

tance  12  inches.  The  headroom  provided  is  at  least  6  feet  verti- 
cally; when  the  pitch  is  less  than  45°  the  rule  obtains  making 
the  rise  plus  the  tread  in  inches  equal  24;  below  20°  pitch  noth- 
ing is  required  except  occasionally  a  plank. 

Though  used  in  Europe  for  1200  to  1500  feet  depth,  and  in 
this  country  in  deep  mines,  ladders  are  certainly  not  ad- 
visable. According  to  the  Cornwall  Society,  their  use  deranges 
the  respiration  and  shortens  life  by  ten  years.  The  miners 
reach  the  workings  more  or  less  exhausted,  and  the  operators 
have  lost  the  benefit  of  a  proportionate  amount  of  energy.  Unques- 
tionably, an  element  of  success  worthy  of  attention  by  mine  mana- 
gers— a  pecuniary  as  much  as  a  humanitarian  question — is  the 
proper  treatment  of  and  the  conveniences  for  the  men,  who 
unconsciously  reciprocate  in  an  equivalent  of  work.  Ladders 
waste  time.  It  takes  15  minutes  to  go  down  300  feet,  and  the 
ascent  is  twice  as  slow.  A  shift  of  forty  men,  following  one 
another  at  intervals  of  8  feet,  entails  a  loss  to  the  company  of  31 
minutes  each  shift.  With  buckets  and  cages  the  loss  is  not  so 
great;  eight  men  at  a  time,  lowered  1200  feet,  consume  40 
minutes  for  every  shift  of  100  men.  An  additional  loss  occurs 
at  tally-time  from  the  reduction  of  the  hoisting  capacity,  which, 
with  the  impatience  of  the  men,  leads  to  the  crowding  of  the 
cage;  but  in  most  States  the  limiting  number  of  men  permitted 
on  the  cage  is  named. 

The  Man-engine. — Movable  ladders  or  man-engines,  invented 
by  D'Orrell,  of  Clausthal,  were  adopted  as  acceptable  substitutes 
for  ladders  and  were  once  popular  in  deep  mines.  When  intro- 
duced into  Cornwall  by  Mr.  Lorn,  the  engine  was  declared  by 
the  Royal  Polytechnic  Society  a  very  tl  great  boon  to  miners." 
Its  introduction  involved  the  addition  of  some  machinery,  and, 
though  easy  to  operate,  it  is  now  obsolete  as  an  economic 
arrangement. 

Two  rods,  of  decreasing  cross-section  from  the  top  down,  receive 
at  the  surface  an  oscillatory  motion  from  balanced  bobs,  operated 
by  an  engine  having  a  fly-wheel  and  other  regulators.  The 
dimensions  of  each  rod  at  any  point  must  be  such  that  it  will  have 


HOISTING  MACHINERY,  ETC. 


253 


the  requisite  tensile  strength  to  support  the  weight  of  the  part 
below  it,  loaded  with  men.  The  rods  play  between  roller-guides 
50  feet  apart,  and  are  provided  with  wings  and  catches,  after  the 
manner  of  the  Cornish  pump-rods,  which  may,  in  fact,  be  utilized 
as"Fahrkunst"  rods  without  much  extra  power. 

Each  rod  has  a  small  platform  (Fig.  no)  about  i2"Xi2"  or 
1 8"  at  every  12  feet — double  the  length  of  the  stroke.  A 
handle  4  feet  above  the  platform  gives  support  to  the  miner,  who 
is  carried  up  6  feet  on  one  rod,  which  brings 
him  opposite  a  platform  on  the  companion 
.  rod;  upon  this  he  steps,  to  be  lifted  6  feet 
more,  to  meet  a  plat  on  the  first  rod,  which 
has  been  coming  down  to  receive  him.  A 
miner  stepping  from  one  to  the  other  is  car- 
ried up  or  down  at  a  rate  of  from  48  to  96  feet 
per  minute  (each  rod  makes  4  to  8  double 
strokes,  delivering  one  man  each  time;  those 
at  the  Calumet  and  Hecla  make  five  strokes). 
As  there  is  no  limit  to  the  depth  to  which  these 
may  be  carried,  and  as  they  are  capable  of 
working  alike  in  slopes  as  in  shafts,  it  is  not 
surprising  that  they  "take"  so  well.  They 
replace  hoisters,  and  require  little  additional 
power  or  space.  Tools  and  supplies  cannot 
be  carried  by  the  miner,  but  may  be  delivered 
by  the  cage  or  bucket. 

A  single  rod  is  also  used,  its  companion 
being  replaced  by  stationary  platforms  at- 
tached, 6  feet  apart,  to  the  shaft-timbers. 
Upon  these  the  ascending  men  wait  during  the 
down-stroke  of  the  rod.  The  single-acting  man-engine  requires 
chains  and  counterpoises  at  intervals,  to  balance  it  and  to  prevent 
the  shock  incurred  at  the  end  of  the  stroke. 

From  the  fact  that  a  misstep  would  be  fatal,  it  would  seem 
as  though  man-engines  were  extra-hazardous,  yet  the  accident 
record  does  not  confirm  this  belief.  Some  confusion  is  caused 


FIG.  no. 


254  MANUAL  OF  MINING. 

by  a  man  missing  his  plat  and  riding  on,  to  the  annoyance  of 
those  following  him;  but  this  is  of  rare  occurrence  unless  his 
light  goes  out,  for  there  is  a  halt  of  several  seconds  at  each  change 
of  motion.  In  Prussia,  out  of  an  average  of  100,000  men  employed 
for  ten  years,  only  57  were  injured  on  the  man-engines;  in 
Cornwall,  17.  This  is  more  than  compensated  for  by  the  increased 
length  of  life  of  the  miners  using  them. 

The  cost  of  machinery,  etc.,  for  a  1200- foot  man-engine  is 
$18,000,  upon  which  interest  and  depreciation  may  be  figured  at 
$2500, — amounting  to  10  cents  per  man  daily  on  a  gang  of  100 
men.  The  running  expenses  at  the  Dolcoath  mine  are  4  cents 
per  man  for  2400  feet. 


REFERENCES. 

Winding  Ropes  in  Belgium,  Coll.  Guard.,  Jan.  16,  1903;  Glossary  of 
Rope  Terms,  Mines  &  Min.,  April  1904;  Notes  on  Steel-wire  Cables,  Prof. 
F.  Soule,  A.  I.  M.  E.,  Vol.  XXIX,  550;  The  Testing  of  Winding  Ropes  in 
the  Province  of  Anhalt,  Germany,  F.  H.  Probert,  A.  I.  M.  E.,  Vol.  XXX, 
1020;  The  Hand  Auger  and  Hand  Drill  in  Prospecting  Work,  C.  Catlett, 
A.  I.  M.  E.,  Vol.  XXVII,  123;  Mine  Ropes:  Strength,  Care,  etc.,  C.  Aquil- 
lon,  Colo.  S.  of  M.  Quart.,  I,  1892,  20;  Vegetable-fiber  Ropes,  M.  &  M., 
Vol.  XXIV,  530;  Testing  Wire  Ropes,  M.  &  M.,  Vol.  XXIV,  405;  Cable 
Wires,  M.  &  M.,  Vol.  XXIV,  411;  Stresses  in  Wire  Ropes,  M.  &  M.,  Vol. 
XXII,  316;  Due  to  Shock,  M.  &  M.,  Vol.  XX;  Mine  Ropes:  Strength,  Care, 
etc.,  Colo.  S.  of  M.  Quart.,  I,  1892,  20;  Hoisting  Ropes,  George  S.  Whyte, 
M.  &  M.,  April  1904;  Splicing  Wire  Ropes,  M.  &  M.,  April  1904. 

Head  Frame  Illustrated,  Penna.  Mine  Insp.  Packet,  1886,  1889,  104; 
1891,  395;  1889,  240;  1890,  389;  Head  Works  Framing,  Charles  H.  Fitch) 
Min.  &  Sci.  Press,  June  13,  1903;  Steel  Head  Frames,  M.  &  M.,  Vol.  XX,' 
292;  and  Steel  for  Coal-mine  Tipples,  Jour.  W.  Soc.  of  Eng.,  IX,  343,  and 
X,  383- 

Combined  Safety  Clutch  and  Stopping  Device  or  Winding  Cages,  O. 
Schenck,  Coll.  Guard.,  April  12,  1900;  Safety  Clutch,  111.  Min.  Inst,  II,  109, 
159,  180,  212;  Safety  Catches,  Automatic  Appliances  in  Winding,  C.  Roemer, 
Coll.  Guard.,  Nov.  1896,  916;  Safety  Clutch,  Mine  Insp.,  Pa.,  1886,  T\b; 
Attachments,  111.  Min.  Inst.,  II,  109,  159,  180,  212. 

Tipping  Cars,  Illustrations,  Coll.  Guard.,  April  30,  1897;  Spring  Coup- 
lings (Cage  to  Rope),  Trans.  M.  &  M.  Eng.,  XLVI,  54;  The  Tipping  and 
Screening  of  Coal,  J.  Rigg,  N.  Staff.,  IV,  103;  Dumping  Cradles  for  Mine 


HOISTING  MACHINERY,  ETC.  255 

Cars,  A.  I.  M.  E.,  XVII,  564;  Coll.  Guard.,  April  30,  1897,  824;  Colliery 
Surface  Works,  E.  B.  Wain,  Coll.  Guard.,  Dec.  1904,  1076;  A  Steel  Coal 
Tipple,  Eng.  News,  June  1904,  577. 

Slope  Carriage,  Pa.  Mine  Insp.,  1882,  39,  18. 

Steel  Mine  Cars,  M.  &  M.,  Vol.  XXII,  i4,:  Steel  Wheels,  M.  &  M.,  Vol. 
XXII,  364;  Cars,  A.  I.  M.  E.,  Vol.  V,  502;  Latches,  Dogs,  etc.,  Hildbrandt, 
essay,  23;  The  Comparative  Merits  of  Large  and  Small  Trams  for  Colliery 
Use,  James  Brogden,  So.  Wales  Inst.,  VI,  173;  Car  Clutches,  Hildbrandt, 
essay,  42;  Wheels,  Self-oiling,  Mine  Insp.,  Pa.,  1889,  no;  Note  on  the 
Friction  of  Car  Wheels,  R.  Van  A.  Norris,  Amer.  Inst.  M.  E.,  XVIII,  508; 
Packing  and  Lubrication,  R.R.  Gaz.,  Jan.  26,  1900;  Car  for  Handling  Rock 
at  Mines,  L.  L.  Logan,  M.  &  M.,  Oct.  1903. 

Ladders,  Strength  of:  Tables,  etc.,  R.  G.  Browning,  Eng.  &  M.  Jour., 
June  1897,  602. 

Mine-car  Wheels,  Coll.  Guard.,  LXXX,  392;  Self -lubricating  Wheels, 
Coll.  Guard.,  Vol.  LXXX,  392;  Automatic  Stop  for  Cars,  Coll.  Guard.,  Vol. 
LXXIX,  66;  Life  of  Cables,  Coll.  Guard.,  Vol.  LXXIX,  224;  Pit  Ponies,  Coll. 
Guard.,  Vol.  LXXIX,  1092;  Detaching  Hooks,  Coll.  Guard.,  Vol.  LXXXI, 
369;  Testing  Hoist  Ropes  in  Germany,  Coll.  Guard.,  Vol.  LXXXI,  456; 
Telescope  Door  for  Cages,  Coll.  Guard.,  Vol.  LXXXII,  415;  Working  Loads 
for  Manilla  Rope,  Coll.  Guard.,  Vol.  LXXXIII,  143;  Belgian  Hoist  Ropes> 
Coll.  Guard.,  Vol.  LXXXI V,  123;  Changing  Winding  Ropes  on  Koepe  Pul- 
leys,  Coll.  Guard.,  Vol.  LXXXVI,  1130,  and  LXXXV,  425;  Safety  Measure* 
against  Overwinding,  Coll.  Guard.,  Vol.  LXXXVI,  1197. 

Detaching  Hooks  for  Cages  and  Ropes,  111.  Min.  Inst.,  II,  109.  159,  180, 
212;  Shaft  Accidents  and  Preventions  at  Level  Landings,  N.  Staff.  Inst., 
VIII,  204;  Fencing  Gates  for  Winding  Shafts,  W.  Hay,  Inst.  M.  E.,  X. 

Overwinding  Precautions,  Coll.  Guard.,  Jan.  22,  1897,  174;  Its  Preven- 
tion at  Collieries  and  Furnaces,  W.  Grimmitt,  Fed.  Inst.  M.  E.,  II;  Its  Pre- 
vention by  a  Patent  Apparatus,  Fed.  Inst.  M.  E.,  I;  Overwinding  Precautions, 
Joel  Settle,  Coll.  Guard.,  Jan.  1897,  174;  Accidents  from  Overwinding  and 
other  Cage  Accidents  in  Mines,  Coll.  Mang.,  May  1893,  88;  Overwinding 
Safety  Devices,  Mining  Bulletin,  P.  S.  C.,  1896;  M.  &  M.,  Vol.  XXII,  295. 

Colliery  Shaft  Signalling,  G.  W.  Smith,  Min.  Inst.,  Scot;  Closing  Top  of 
Upcasts,  A.  Reid,  Fed.  Inst.  M.  E.,  X;  Electric  Signals  for  Collieries,  C. 
McLaren  Irvine,  Min.  Inst.,  Scot.,  IV,  47;  Cages,  Illustrated,  Pa.  Mine 
Insp.,  1884,  563;  Mallisard-Taza  Tipping  Cage  used  in  French  Collieries, 
Eng.  &  Min.  Jour.,  I,  129;  Stops  for  Mine  Cages,  Coll.  Guard.,  1896,  1173, 
916;  Signalling  between  Levels,  F.  B.  Allen,  N.  Z.  Mines  Rec.,  Aug.  17, 
1903;  Hoisting  Ropes,  Robert  Peele,  Mines  &  Min.,  Mar.  1900;  Mine  Sig- 
naling by  Compressed  Air,  Bernard  McDonald  and  William  Thompson, 
Con.  Min.  Rev.,  Sept.  30,  1002;  Signal  Device  for  Mines,  C.  S.  Hertzig, 


256  MANUAL  OF  MINING. 

A.  I.  M.  E.,  Vol.  XXX,  314;  Electric  Mining  Signals,  M.  &  M.,  Vol.  XXII, 
167. 

Kips  or  Landing  at  Shafts,  M.  H.  Douglass,  Brit.  Soc.  Min.  Stud.,  I,  433; 
Shaft  Timbering,  J.  C.  Jefferson,  Chest.  Inst.,  VIII,  209;  Some  Arrange- 
ments for  Preventing  Accidents  at  Level  Landings  in  Cage  Dips  and  Shafts, 
A.  R.  Sawyer,  N.  Staff.  Inst.,  VIII,  204. 


CHAPTER  VIII. 

UNDERGROUND  HAULAGE  SYSTEMS. 

The  Extent  of  Haulage  Requirements. — In  any  coal-mine  of 
moderate  size  there  are  fully  500  miles  of  track  and  headings 
into  the  rooms,  representing  perhaps  2000  miles  of  road,  which 
at  some  time  during  its  history  had  been  laid  down  and  given 
service.  It  has  been  estimated  that  in  the  production  of  200,- 
000,000  tons  of  coal  at  least  5000  miles  of  track  extension  are 
made  each  year.  The  great  amount  of  it  indicates  the  necessity 
for  giving  special  attention  to  its  construction.  The  cost  of 
delivering  coal  from  the  working  face  to  the  tipple  when  the 
tracks  are  level  and  the  distance  not  over  a  mile  is  between  6 
and  10  cents  per  ton.  During  a  single  month  an  average  of 
thirty-one  mines  using  rope  haulage  showed  a  cost  of  7.5  cents 
per  ton.  In  twenty  mines  employing  electric  haulage  the  run- 
of-mine  coal  cost  6  cents  per  ton,  and  8  cents  in  an  average 
of  twenty-one  other  mines  having  mule  haulage.  Probably, 
over  the  United  States,  the  average  cost  of  animal  haul  is  nearly 
15  cents.  The  cost  of  haulage  increases  materially  as  the  distance 
of  travel  exceeds  3000  feet.  Beyond  this  length  of  haul  it  may 
be  advisable  to  sink  another  shaft  if  a  convenient  point  of  ship- 
ment can  be  obtained  at  the  surface  and  the  depth  of  shaft  be 
not  too  great. 

The  Grade  and  Direction  of  a  Haulage  Road. — A  haulage 
road  is  as  straight  and  direct  as  circumstances  will  allow,  so 
designed  as  to  utilize  the  natural  forces  to  the  closest  degree. 
The  shaft  and  slope  are  carried  down  to  the  lowest  point,  that 
to  it  may  be  delivered  the  mineral  from  the  working  face  on  a 

257 


258  MANUAL  OF  MINING. 

continual  descent.  Dropped  through  the  mill-holes  from  stope 
to  cars,  the  ore  is  moved  by  hand-  or  horse-power  to  the  foot 
of  the  shaft;  and  coal,  leaving  the  face  in  cars,  should  have 
the  advantage  of  the  down-grades  in  its  delivery  to  the  shaft 
bottom. 

From  the  workings  to  the  haulage-way  the  inclination  of  the 
track  will  depend  solely  upon  the  mean  inclination  of  the  floor, 
but  the  cars  should  be  capable  of  moving  by  gravity  without 
the  use  of  blocks  or  sprags.  A  sprag  is  a  billet  of  wood 
inserted  between  the  spokes  to  prevent  the  wheels  from  revolv- 
ing, thus  converting  rolling  friction  into  sliding  friction  and 
checking  the  speed  of  the  car.  On  a  3  per  cent  grade  only  one 
sprag  is  used,  but  on  a  10  per  cent  grade  four  sprags  will  be 
necessary  to  keep  the  speed  of  the  car  within  moderate  control. 
A  grade  exceeding  one  in  five  is  positively  dangerous.  Roads 
•are  often  designated  as  one-,  two-,  three-,  or  four-sprag  roads, 
according  to  the  number  of  wheels  which  must  be  blocked. 

The  best  grade  for  any  haulage  road  is  the  gradient  of  equal 
resistance — that  on  which  the  work  expended  in  hauling  a  loaded 
car  down  equals  that  of  the  returning  car  up.  A  steeper  grade 
than  this  is  wasteful  of  power  both  ways.  As  evidencing  the  in- 
fluence of  grade  in  the  reduction  of  useful  power,  it  may  be  cited 
that  a  locomotive  can  pull  up  an  incline  of  2  feet  in  a  hundred 
only  80  per  cent  of  what  it  can  haul  on  a  level,  and  on  a  4  per  cent 
grade  but  30  per  cent.  This  is  correspondingly  true  of  animal 
haulage.  While  a  mule  can  haul  three  loaded  cars  on  a  reason- 
ably level  track,  it  can  haul  only  one  on  a  6  per  cent  grade. 

Haulage-ways  in  Metal-mines. — Metalliferous  mines  rarely 
present  any  difficulty  for  the  road  engineer.  Very  few  roadways 
feed  to  any  given  landing.  In  a  vein  there  is  but  one  line  on 
either  side  of  the  shaft  for  which  provision  is  to  be  made.  In  large 
ore  bodies  only  a  few  tracks  will  radiate  from  the  point  of  outlet, 
and  but  few  cars  will  travel  over  each. 

Any  desired  grade  may  be  laid  in  vein  mines.  It  is  as  much 
greater  than  that  required  for  drainage  as  the  available  motor 
power  will  allow  for  returning  with  the  empty  cars,  or  controlling 


UNDERGROUND   HAULAGE  SYSTEMS.  259 

the  speed  of  the  loaded  cars  out.  The  direction  of  the  roadway 
is  fixed  by  the  sinuosities  of  the  vein  wall.  For  animals  due 
regard  must  be  had  to  economy  of  size  and  weight  of  car  and 
its  contents;  hence  the  grade  will  probably  not  exceed  1.5  in 
100  feet.  For  locomotive  haulage  the  grade  may  be  carried  to 
4  per  cent  as  a  maximum. 

Haulage-ways  in  Coal-mines. — Here  the  engineer  must  accom- 
modate himself  to  the  stratigraphical  conditions.  The  thickness 
of  the  coal-bed  and  the  frequency  and  amount  of  the  existing 
faults,  folds,  and  inclinations  must  be  reckoned  with,  and  a  uni- 
form grading  would  demand  their  elimination,  according  to  the 
character  of  the  enclosing  rock. 

In  the  main  headings  the  hillocks  may  be  removed  and  the 
sumps  in  the  floor  filled,  or  the  roof  material  over  the  depressions 
may  be  ripped  down  and  the  sumps  filled  with  the  waste.  The 
latter  is  more  economical.  It  is  far  more  expensive  to  take  up 
a  foot  of  bottom  than  to  blast  down  3  feet  of  thickness  in  the  roof. 
Again,  the  cost  of  maintenance  is  not  so  great  where  the  road-bed 
filling  is  of  loose  slate  from  the  roof,  as  in  the  customary  floor  of 
fire-clay  or  limestone.  Often  the  soft  floor  of  the  coal-seams 
cannot  be  disturbed,  and  it  may  not  be  desirable  either  to  fill 
up  the  depressions  or  to  remove  the  hills,  because  of  inconvenience 
in  making  connections  to  rooms.  It  is  probable,  too,  that  the 
cleavage  planes  in  the  coal  will  determine  the  directions  of  the 
haulage-ways,  and  the  average  dip  of  the  vein  will  determine 
their  grade. 

In  coal-mines  there  are  always  two  parallel  headings  for 
ventilation  purposes,  the  engineer  having  his  choice  as  to 
which  shall  drain  the  water  and  which  be  used  for  haulage. 
It  is  rare  that  the  same  heading  would  be  required  to  give 
both  services.  Usually  the  intake  is  the  one  employed  for 
haulage,  because  of  the  better  quality  of  air  and,  in  consequence, 
a  less  cost  of  maintaining  the  track.  This  roadway  is  never 
obstructed  by  the  air-crossings,  which  usually  are  placed  on  the 
return-airway.  When  no  other  condition  determines  the  choice, 
the  roadway  higher  in  elevation  is  employed  for  haulage  and 


260  MANUAL   OF  MINING. 

the  lower  one  for  drainage,  with  frequent  communications  be- 
tween them  to  allow  of  the  accumulated  waters  of  the  former  to 
be  promptly  carried  to  the  latter. 

The  haulage  gangway  is  of  single-track  width  only  when  the 
system  employed  is  the  tail-rope  or  the  locomotive.  The  double- 
track  gangway  is  employed  where  the  traffic  is  heavy,  or  for  an 
endless-rope  haulage-way.  The  advantage  in  the  latter  is  the 
great  area  given  to  the  air-current,  and  also  the  opportunity 
afforded  for  utilizing  the  excess  of  the  power  in  the  down-going 
car  or  train  to  assist  the  engine  in  raising  the  load  on  the  other 
track. 

The  requirements  for  ventilation  demand  a  roadway  of  liberal 
area  which  will  permit  a  large  quantity  of  air  to  pass  through  with 
the  minimum  of  friction,  and  also  reduce  the  effect  of  the  dis- 
turbance to  the  current  by  the  passage  of  the  train  through  it. 

The  enlargement  of  the  area  necessary  to  provide  for  the  cars 
may  be  obtained  by  increasing  the  width  or  the  height  of  the 
roadway.  The  former  is  preferable  unless  the  roof  is  very  firm 
and  the  bed  quite  thick.  The  objection  to  it,  however,  is  the 
increased  length  of  sills  and  the  timbers  necessary  to  maintain  a 
good  level  on  the  ordinary  floor  of  the  coal-mine.  It  is  prefer- 
able to  increase  the  height  unless  the  roof  will  not  bear  cutting 
down.  It  is  a  cheaper  method,  for  the  dimensions  of  the  sills 
would  be  only  of  moderate  size,  and  those  of  the  posts  would  not 
be  any  greater  than  for  a  wide  roadway. 

Track  Construction. — The  tracks  on  the  haulage-ways  are 
constructed  with  a  view  not  only  to  diminish  the  working  ex- 
penses, but  also  to  maintain  a  large  capacity  with  a  given  amount 
of  power.  The  difference  in  cost  and  the  time  of  laying  between 
a  heavy  rail  on  a  substantial  track  and  one  which  is  poorly 
laid  is  not  large,  nor  is  it  commensurate  with  the  difference  in 
iheir  efficiencies.  A  reduction  in  the  grade  of  \  per  cent  saves 
annually  in  track  maintenance  an  equivalent  of  30  tons  of  fuel 
for  motor  power  for  each  car  running  through  the  mine.  The  cost 
of  such  construction  might,  however,  prove  prohibitive  of  any 
operations.  The  most  that  can  be  done  is  to  improve  the  roads 


UNDERGROUND  HAULAGE  SYSTEMS.  261 

within  reasonable  limits  of  cost  after  careful  surveys  and  level- 
lings  of  all  roads  intended  for  main  arteries. 

The  road-bed  is  made  as  elastic  as  possible  by  the  use  of  sand, 
dump-coal,  or  ashes,  without  being  too  soft  for  support.  The 
choice  of  ballast  depends  upon  the  character  of  the  mine  floor. 
If  the  latter  is  inclined  to  be  wet  or  rough,  then  rocky  material 
would  be  preferable  to  sand  or  ashes.  When  the  floor  is  very 
strong,  clay  can  be  used  to  advantage  with  fragments  of  rock 
grounded  in.  The  dry  road-bed  can  be  secured  by  rounding 
off  the  surface  between  the  tracks  to  a  slight  arch.  This  will 
assist  the  removal  of  water  from  the  track  to  the  drains. 

The  ties  are  laid  2  feet  apart  between  centres  for  rails  of  15 
to  30  feet  in  length.  To  these  the  rails  are  spiked,  four  spikes 
being  used  on  each  tie.  Care  is  taken  that  they  are  not  all  driven 
near  the  centre  of  the  tie,  causing  it  to  split.  For  all  rails  of 
16  to  20  Ibs.  per  yard  the  ties  are  5  inches  on  the  face,  4  inches 
deep,  and  5^  feet  long.  For  the  heavier  rails  in  the  mine  they 
are  6"X$"  deep.  They  are  laid  normally  to  track,  and  on 
curves  are  laid  in  the  direction  of  the  radii  of  the  curve.  In 
some  mines  the  ties  are  creosoted  to  increase  their  durability. 

The  rails  are  the  common  steel  T  rails,  having  the  same  speci- 
fications as  are  adopted  for  railroad  service.  Their  weight 
varies  from  12  Ibs.  per  yard  in  the  rooms  to  40  Ibs.  per  yard  in 
the  slopes.  They  cannot  be  too  heavy  within  reasonable  limits. 
The  greater  the  pitch  of  the  slope  and  the  greater  the  speed  of 
haulage,  the  larger  and  heavier  should  be  the  construction  of  the 
track.  The  rail  for  locomotive  haulage-ways  requires  a  weight 
of  10  Ibs.  per  yard  for  each  ton  that  falls  upon  the  drivers.  The 
cost  of  one  mile  of  i6-lb.  rails  is  about  $1600. 

The  track  in  the  room  and  at  the  long  wall-face  is  laid  as 
rapidly  as  the  mining  operations  progress,  and  with  as  great  care 
to  secure  a  firm  run  as  in  the  main  ways.  It  is  often  laid  in  sec- 
tions composed  of  two  rails  of  15  feet  each.  These  sections  are 
portable  and  carried  up  to  the  face  of  the  workings  with  the 
progress  of  the  room,  the  expense  of  such  track-laying  usually 
being  included  in  the  price  paid  to  the  miner  per  ton  of  coal. 


262  MANUAL  OF  MINING. 

Their  advantage  lies  in  their  simplicity,  strength,  and  durability. 
The  gauge  does  not  depend  on  the  skill  of  the  workman,  and 
will  be  preserved  as  long  as  the  line  lasts.  The  lightest  rail 
advisable  for  such  purposes  is  that  which  weighs  16  Ibs.  to  the 
yard.  Wooden  rails  are  no  longer  employed  even  for  tempo- 
rary service  in  the  rooms. 

The  Gauge  of  Track  is  rarely  less  than  30  inches  or  more 
than  44  inches.  Local  conditions  determine  the  choice.  A 
broad  gauge  gives  a  greater  stability  to  the  cars  and  a  reduction 
of  haulage  expenses,  but  is  more  costly  to  build.  The  minimum 
gauge  of  30  inches  affords  easy  haulage,  opportunity  for  sharp 
curves,  and  a  cheaper  track. 

Turnouts. — The  sidings  are  formed  by  a  set  of  points  and 
crossings,  either  right-  or  left-hand,  a  section  of  curved  line,  and 
a  number  of  sections  of  straight  line. 

Switches  from  one  track  to  another  are  obtained  by  the  use 
of  movable  points  similar  to  those  employed  on  the  surface 
roads,  though  shorter  in  length.  The  point  is  simply  a  spike 
driven  into  one  of  the  ties  supporting  the  rails.  It  is  held  in  posi- 
tion by  and  turned  upon  an  iron  pin  passing  through  the  plate 
welded  to  the  base  of  this  point  and  thrown  out  of  line  to  the 
branch  side  when  the  car  is  to  be  switched.  When  two  of  these 
points  are  connected  by  a  flat  piece  of  iron  passing  under  the 
rails  of  the  track,  the  pair  can  readily  be  thrown  right  or  left  by 
a  bell-crank  lever,  and  both  rails  of  the  track  are  opened  for  the 
switch.  Wherever  possible  single  sets  of  points  are  arranged 
for,  the  wear  being  less  on  two  pairs  of  single  points  than  on  one 
set  of  double  points,  besides  which  single  sets  are  simpler  to  lay 
down  and  keep  in  order  The  use  of  the  three-way  points  and 
crossings  for  two  opposite  side  branches  are  avoided  underground 
by  turning  the  right-  and  left-hand  lines  from  the  straight  track 
in  positions  which  are  not  quite  opposite  to  each  other.  The 
points  and  crossings  are  made  of  standard  sizes  for  curves  of  any 
special  radius  required  to  suit  the  gauge  of  the  rails  and  the  wheel- 
base  of  the  engines.  All  frogs  and  switches  at  the  shaft-landings 
or  slope  bottom  are  of  pointed  sections  of  standard  rail  sizes, 


UNDERGROUND  HAULAGE  SYSTEMS.  263 

manipulated  by  a  spring  or  a  hand-lever.  For  locomotives  on  a 
main  tramway  the  curves  should  not,  as  a  rule,  have  a  radius 
less  than  from  100  to  125  feet. 

When  one  rail  intersects  another,  frog  crossings  are  used. 
They  are  either  plain  rail  intersections,  broken  at  the  points  of 
crossing,  or  they  are  the  standard  frogs  furnished  by  manufac- 
turers. The  points,  rods,  and  levers  used  underground  are  usu- 
ally made  and  fitted  by  the  colliery  blacksmith,  the  plate-layer 
assisting  in  the  fitting. 

In  passing  from  a  heading  or  entry  to  the  room  a  "parting " 
is  used.  This  is  similar  to  a  switch,  the  wings  running  back- 
ward from  the  frog  end  near  the  rails,  but  are  cut  in  a  slanting 
direction  in  order  to  provide  a  sufficient  clearance  between  the 
point  and  the  rail  to  allow  the  wheels  of  the  car  to  pass  along 
the  main  track  when  it  is  not  to  be  deflected  into  the  room.  There 
are  two  forms  of  partings,  one  of  rails  and  another  a  combination 
of  rails  and  cast-iron  pieces,  the  latter  being  much  inferior  to 
the  former.  The  former  has  less  pieces,  gives  more  solid  track, 
and  is  less  liable  to  derail  the  cars.  It  is  cheaper  and  can  be  made 
in  the  blacksmith-shop. 

Turnouts  are  obtained  by  the  use  of  plates  or  turntables. 
Of  the  former  the  simple  iron  plates  laid  on  stout  planks 
revolving  around  a  vertical  pivot  are  employed  where  animal 
power  is  used.  These  are  common  in  metalliferous  mines, 
being  preferred  to  self-acting  switches.  Not  infrequently  is 
seen  at  intersections  of  roads  a  large  fixed  iron  plate,  raised  upon 
suitable  supports,  level  with  the  road-bed,  and  having  curved 
guide-rails  at  each  of  the  four  corners.  These  may  be  em- 
ployed as  turntables  by  dragging  the  car  around  on  the  plate 
until  it  faces  the  desired  track,  when  it  is  run  off.  This  crude 
device  serves  well,  but  has  nothing  to  commend  it. 

Curves. — Curves  in  the  track  are  built  in  the  same  way  as 
on  the  surface,  the  radius  of  curvature,  however,  being  very 
much  shorter.  Connecting  two  entries  on  the  main  haulage- 
road,  the  minimum  radius  for  narrow-gauge  tracks  is  60  feet,  and 
for  a  gauge  of  45  inches  100  feet.  At  the  shaft-landings  the 


264  MANUAL  OF  MINING. 

radius  is  as  short  as  30  feet  for  36-inch  gauge,  and  50  feet  for 
44-inch  gauge. 

The  Degree  of  Curvature  of  a  Curve. — By  this  term  is  under- 
stood the  number  of  degrees  of  circular  arc  measured  at  the 
centre  which  will  be  subtended  by  a  chord  of  100  feet  length. 
Thus  a  10°  curve  is  one  in  which  a  loo-foot  chord  at  the  circum- 
ference is  comprised  between  two  radii  making  25°  at  the  center. 
The  radius  of  a  one-degree  curve  is  practically  5730  feet.  Hence 
the  10°  curve  will  have  a  radius  of  573.0  feet.  This  is  only 
approximately  correct,  but  is  sufficiently  accurate  as  a  standard  of 
measure  for  engineers. 

All  curves  on  underground  roads  are  laid  by  transit  with 
care  equal  to  that  employed  on  the  surface  for  larger  work.  The 
central  line  between  the  rails  of  each  track  is  staked  out  at  inter- 
vals of  20  feet,  from  which  on  the  other  side  are  laid  down  the 
two  rails.  The  latter  are  now  bent  to  conform  to  the  curvature 
of  the  circle  on  which  the  rails  are  laid. 

Elevation  of  the  Outer  Rail. — Provision  is  made  for  neutral- 
izing the  effect  of  centrifugal  force,  or  the  pull  of  the  rope,  on  the 
curve  by  elevating  one  rail  an  amount  depending  upon  its  sharp- 
ness. On  curves  used  by  locomotives  the  outer  rail  is  the  one 
which  is  elevated,  as  also  on  the  slope  haulage-ways  when  the 
weight  of  the  cars  draws  the  main  rope  off  its  drum.  In  a  slope 
haulage  the  outer  rail  is  raised  i  inch  on  a  9°  curve,  2  inches  on 
an  1 8°  curve,  and  4  inches  on  a  36°  curve,  the  latter  being  the 
maximum.  It  never  exceeds  4^  inches  in  height  on  a  curve  of 
50  feet  radius,  or  of  114°  curvature,  when  the  gauge'is  42  inches. 
This  would  be  suited  to  the  ordinary  speed  of  locomotive  haulage, 
or  that  of  cars  running  down  the  slope  by  gravity. 

The  inner  rail  is  elevated  above  the  outer  one  only  on  such 
curves  and  under  such  conditions  as  those  in  which  the  action 
of  the  centrifugal  force  is  less  than  that  of  the  pull  of  the  deflected 
rope.  In  this  case  the  rope  passes  over  rollers  near  the  inner 
rail  and  tends  to  pull  the  car  inward.  It  is  possible,  by  comparing 
the  value  for  centrifugal  force  under  given  conditions,  as  cal- 
culated below,  with  that  of  the  pull  of  the  rope  upon  the  car,  to 


UNDERGROUND   HAULAGE  SYSTEMS.  265 

determine  the  amount  of  the  inclination  which  must  be  given 
to  either  the  inner  or  the  outer  rail. 
Let  g=  gauge  of  the  track  in  inches; 

e  =  elevation  of  the  outer  rail  in  inches; 

R=  radius  of  track  in  feet; 

v=  velocity  in  feet  per  second; 

D  =  degree  of  curvature  of  centre  of  track; 

c=  centrifugal  force  due  to  the  speed  around  the  curve. 

Then      e=°'°3lgV  =o.ooooo$4V2Dg. 

K 

Conveniences  at  Landings. — Whatever  may  be  the  system  of 
haulage,  arrangements  are  made  for  the  prompt  handling  of 
cars  at  the  shaft  bottom.  When  the  coal  is  hauled  from  the  dip 
by  rope  or  locomotive,  the  slope  is  continued,  if  possible,  on  the 
same  side  of  the  hoisting-shaft  at  a  convenient  distance  from  it, 
where  the  requisite  number  of  loaded  cars  can  be  accumulated 
and  allowed  to  run  by  gravity  on  a  one  per  cent  grade  to  the 
shaft. 

The  empty  car  can  be  run  off  the  cage  on  to  a  short  down- 
grade of  about  4  per  cent,  and,  with  the  momentum  given  it  by 
the  loaded  car  being  run  on  the  cage,  would  descend  far  enough 
to  rise  up  on  an  opposing  grade  of  about  2  per  cent  to  an  auto- 
matic switch,  which  delivers  it  to  the  empty-car  siding.  Here 
it  may  be  taken  up  by  the  locomotive,  or  a  rope  passing  around 
the  shaft  through  an  entry  designed  for  this  purpose.  Some- 
times a  short  chain  haulage  operated  by  an  electric  motor  is 
employed  to  draw  the  empty  cars  to  the  main  haulage-way  when 
the  gravity  system  cannot  be  employed. 

The  landings  for  the  cars  in  the  mine  at  the  end  of  the  end- 
less-rope haulage-way  are  reached  by  switches  from  the  main 
track  by  slide-knuckles,  a  place  being  allowed  for  the  rope  to 
pass  through  the  switch  without  injury  from  the  flanges  of  the 
wheels  passing  over  it. 

Safety  Devices  Along  Haulage-ways. — To  prevent  accident 
from  runaway  cars  on  a  grade,  whether  underground  or  at  the 
surface,  automatic  devices  are  employed  either  to  check  the  speed 


266  MANUAL  OF  MINING. 

of  the  car  or  to  automatically  shunt  it  off  the  track.  Cars  are 
frequently  fitted  with  a  drag  projecting  below  the  car,  which,  when 
the  latter  breaks  away  from  the  hoist-rope,  catches  in  the  floor  of 
the  track  and  prevents  the  runaway.  A  similar  device  is  employed 
whereby  a  car  trips  a  block  which  is  thrown  across  the  track 
before  the  car  reaches  it  and  stops  the  car,  preventing  further 
damage  being  done  by  its  descent  to  the  lower  end  of  the  plane. 

Catches  at  the  Top  of  the  Incline. — These  are  to  prevent  cars 
from  breaking  away  at  the  top  of  the  plane.  Reaumaux's  catch 
consists  of  a  bent  lever  with  two  unequal  arms  turning  around  a 
vertical  axis.  It  is  placed  flat  on  the  ground  and  is  parallel  to 
the  tramway  and  at  such  a  distance  that  at  least  one  of  the  ends 
always  bars  it,  with  the  short  arm  next  the  side  of  the  inclined 
plane.  As  the  empty  car  reaches  the  top  of  the  incline  and  passes 
forward  its  front  wheel  touches  the  long  arm  after  the  hind  wheel 
has  passed  the  short  arm,  and  the  way  is  blocked.  Another  run 
on  the  same  track  cannot  be  made  until  the  protruding  arm  is 
turned  back  by  the  planeman. 

Blocks  or  stops  are  simpler  appliances.  There  are  two  pieces 
of  wood  at  right  angles  to  one  another  moving  on  an  upright  pin. 
One  arm  is  thrown  across  the  track  to  hold  the  car  in  place  until 
the  latter  is  to  be  released  and  lowered  (6,  Fig.  84).  Then  it  is 
knocked  out.  The  same  plan  is  adopted  in  10,  Fig.  84,  by  the 
weighted  lever. 

Stops  Along  the  Incline. — To  guard  against  accidents  from 
runaways  along  the  inclines  a  heavy  iron  bar  is  swung  as  shown 
in  9,  Fig.  84.  It  interposes  no  obstruction  on  the  track  until 
some  one  pulls  the  wire  and  releases  it,  as  shown  by  the  dotted 
lines.  A  similar  balanced  block,  Fig.  84,  is  used,  the  car  being 
caught  by  the  end,  b. 

The  Mortier  safety- catch  consists  of  a  movable  axle  with 
levers  placed  in  the  axis  of  the  roadway  and  supported  upon 
sleepers,  8,  Fig.  84.  In  one  extreme  position  it  is  opened  by 
the  axle  of  the  rising  car;  in  the  other,  it  closes  after  the  train 
passes.  In  the  middle  position  both  levers  bar  the  passage 
The  planeman  can  also  move  the  catch  by  a  pedal. 


UNDERGROUND   HAULAGE  SYSTEMS.  267 

Catches  at  the  Bottom  of  the  Incline. — At  the  bottom  of  the 
incline  are  additional  safeguards.  A  "seizer,"  shown  in  12, 
Fig.  60,  is  placed  at  the  foot  of  the  incline.  An  iron  bent  lever- 
arm  supported  on  a  horizontal  axis  at  the  apex  projects  vertically 
above  the  rails  between  which  the  frame  stands.  A  strip  of  iron 
swinging  horizontally  keeps  up  one  arm,  which  is  inserted  into 
the  link  of  the  chain.  Even  if  the  top  man  should  push  his  train 
over  the  plane  before  the  bottom  train  is  ready,  the  chain  would 
not  move  until  the  upheld  arm  is  released  and  allowed  to  fall 
into  a  recess,  and  thus  free  the  cars.  Various  other  provisions 
of  a  similar  nature  guard  against  the  ill  effects  of  a  broken  end- 
less haulage-chain. 

The  Choice  of  Haulage  Systems. — The  choice  of  a  system 
depends  upon  the  grades  of  the  haulage-ways.  When  the  latter 
are  horizontal,  power  will  be  required  in  both  directions,  and 
obtained  from  the  horse,  locomotive,  stationary  engine,  or  rope 
haul.  When  the  grade  is  toward  the  shaft,  the  loaded  train  may 
be  pulled  with  any  of  these  powers  The  horse  or  mule  is, 
however,  used  to  advantage  only  on  a  favorable  track  of  less  than 
2.5  per  cent.  The  locomotive  may  be  employed  for  grades  not 
exceeding  4  per  cent.  Above  that  it  is  impossible  for  it  either  to 
pull  a  long  train  up  the  inclination  or  to  check  its  descent.  The 
endless-rope  system  or  the  tail-rope  method  is  used  when  the 
grade  does  not  exceed  20  per  cent.  Beyond  this  inclination  the 
plant  will  be  self-acting,  for  the  loaded  train  going  toward  the 
shaft  pulls  the  empty  cars  back  into  the  mine.  When  the  grade 
of  the  haulage-way  is  slightly  against  the  loaded  car,  the  horse 
may  do  for  a  short  haul.  If  it  exceeds  2.5  per  cent,  the  locomotive 
can  no  longer  be  used  advantageously.  On  a  grade  exceeding 
5  per  cent  a  traction  locomotive  is  totally  incapable  of  pulling  any 
load.  Up  to  a  limit  of  10  per  cent,  a  track  locomotive  may  be 
employed.  For  all  steep  inclines  some  rope  system  must  be 
employed. 

Tramming. — Man-power  is  employed  for  incidental  work  on 
the  short  hauls  on  level  ground,  or  in  rooms.  The  weight  of  car 
and  degree  of  inclination  for  such  power  are  necessarily  very 


268  MANUAL  OF  MINING. 

small.  The  average  man  is  capable  of  exerting  a  push  of  27  Ibs. 
at  the  rate  of  2  feet  per  second.  This  would  move  a  2-ton  car  on 
a  level  at  a  speed  capable  of  making  from  three  to  six  round 
trips  of  a  ton  per  mile.  The  condition  of  the  roads  and  running 
gear  will  definitely  fix  this. 

Animal  Haulage. — Horses  and  mules  are  employed  under- 
ground for  limited  purposes  of  transportation  singly  or  tandem, 
according  to  the  length  of  the  trip  and  the  weight  of  the  train. 
It  is  customary  to  work  them  in  traces;  though  in  some  mines 
shafts  are  used  which  give  considerable  holding-back  power  not 
to  be  had  from  traces.  There  can  be  no  objection  to  traces 
on  level  or  nearly  level  roadways.  But  with  loaded  cars  down- 
hill sprags  must  be  placed  in  the  wheels  so  as  to  act  as  a  drag 
for  trace- mules.  The  inclination  of  the  roadway  for  animal 
haulage  power  is  not  over  1.5  per  cent,  as  the  mules  rarely 
take  the  cars  to  the  surface  unless  the  inclination  of  the  entries 
especially  permits  it. 

The  utility  of  the  animal  is  confined  to  haulage  in  the  secondary 
ways,  to  the  rooms,  and  to  switching  where  economy  in  height  must 
be  practised.  Here  the  mule  has  the  superiority  over  the  horse 
because  of  its  shorter  stature  with  equal  strength.  A  thickness 
of  coal-seam  of  3.5  feet  will  accommodate  the  mule,  while  one 
of  4.8  feet  is  essential  for  the  horse.  The  kind  and  size  of  horse 
purchased  for  colliery  working  will  of  course  depend  upon  the 
specific  work  for  which  it  is  required  and  the  nature  of  the  road- 
ways, size  of  cars,  etc.  Before  being  introduced  into  the  pit  they 
are  worked  on  the  surface  for  three  or  four  weeks  to  reduce  the 
risk  of  importing  infectious  diseases  into  the  underground  stud. 
The  possibility  of  new  horses  being  up  to  the  work  expected  of 
them  may  thus  be  ascertained. 

The  ordinary  speed  of  the  animal  is  two  miles  per  hour.  It 
has  a  tractive  force  of  about  150  Ibs.,  at  which  a  single  horse 
or  mule  may  be  estimated  as  an  equivalent  to  from  8  to  9  gross 
tons  of  load  per  hour,  averaging  four  round  trips  of  a  mile  each 
per  day.  The  condition  of  the  track,  the  ventilation,  and  the 
amount  of  delay  at  the  termini  of  the  trip  would  reduce  this 


UNDERGROUND  HAULAGE  SYSTEMS.  269 

average  and  make  the  daily  capacity  of  the  horse  or  mule  about 
4.0  tons.  This  represents  a  gross  yearly  capacity  of  9000  ton- 
miles.  Under  ordinary  conditions  one  mule  will  serve  for  the 
haulage  of  the  output  of  ten  miners.  A  mine  of  1000  tons  daily 
output  will  require  90  mules,  which  with  equipment  will  cost 
$8000. 

The  average  cost  of  mules,  as  well  as  their  keep,  is  nearly 
the  same  as  for  horses.  Their  care  averages  60  cents  per  day, 
inclusive  of  feed.  The  daily  allowance  of  food  varies  with  the 
work  done.  It  is  not  far  from  18  Ibs.  of  grain  and  12  Ibs.  of 
hay  per  animal.  A  less  economical  fodder  now  used  is  "  ensilage." 
This  is  fodder  cut  in  the  field  in  the  usual  way,  but,  instead  of 
being  dried,  is  chopped  up  small  and  stored  green  in  brick  or 
cement-lined  pits,  or  silos.  Dry  food  is  a  more  convenient  feed, 
but  dearer  than  ensilage. 

The  number  of  horses  under  one  ostler  or  keeper  should 
not  exceed  ten.  They  are  generally  stabled  underground,  and 
many,  once  below,  never  see  daylight  again.  Where  the  work- 
ings are  at  a  considerable  distance  from  the  shaft,  the  stables 
are  built  well  back  in  the  interior  of  the  mine,  to  avoid  unneces- 
sary travelling.  The  cost  of  a  comfortable  stable  is  about 
Sio  per  stall.  Special  attention  must  be  given  to  proper  venti- 
lation, drainage,  and  cleanliness.  The  opening  must  be  not 
less  than  18  feet  wide. 

Animal  haulage  is  expensive  and  is  fast  being  supplanted 
by  mechanical  systems.  The  number  of  working  days  at  a 
colliery  seldom  exceeds  200  in  a  year.  The  cost  of  maintaining 
the  animal  during  the  remaining  165  days,  besides  those  result- 
ing from  depression  in  trade,  strikes,  accidents,  and  diseases, 
adds  materially  to  the  cost  of  production.  Moreover,  horses  are 
subject  to  epidemics,  and  it  may  happen  that  in  times  of  greatest 
prosperity  some  or  all  of  the  horses  may  be  unable  to  leave  the 
stables.  Delays  arising  from  the  dropping  of  shoes  while  at 
work  are  costly,  for,  while  the  farrier  is  at  work,  both  horse  and 
driver-  are  standing  idle.  Out  of  every  thirteen  animals  in  the 
mine  an  average  of  ten  may  be  regarded  as  a  full  working 


270  MANUAL  OF  MINING. 

complement,  the  remainder  being  on  the  retired  list  for  various 
causes. 

Compared  with  the  mechanical  methods,  animal  haulage 
is  more  expensive.  It  is  slower  and  therefore  used  on  long  hauls 
only  when  numerous  ventilating-doors  intervene  along  the  line. 

The  Locomotive. — For  great  distances  and  a  large  output 
the  locomotive  is  much  more  efficient,  more  rapid,  and  cheaper 
than  animal  power,  being  equally  flexible  and  having,  moreover, 
the  advantage  in  that  during  strikes  and  lockouts  there  is  no 
expense  in  maintaining  it.  It  can  be  accommodated  to  varying 
demands  of  the  different  sections  of  the  mine.  The  motor  is  in 
direct  charge  of  the  engineer,  who  is  always  on  hand,  so  that 
it  can  be  stopped  promptly.  This  would  avert  such  accidents 
as  are  common  in  the  rope-haulage  systems.  Here  the  con- 
ductor of  the  train  must  first  signal  to  the  surface  before  the 
train  can  be  stopped,  and  any  incumbrances  on  the  track,  or 
the  jumping  of  a  car  from  the  track,  would  result  in  demolition 
of  the  entire  train  before  the  rope  could  be  brought  to  rest. 

The  locomotive  is  operated  by  air,  steam,  or  electricity  with 
nearly  equal  efficiency  and  in  any  desired  size  for  traction  pur- 
poses. It  is  difficult  to  estimate  the  number  in  use  in  the  United 
States,  but  in  Pennsylvania  alone  are  950  locomotives,  503  being 
steam-locomotives,  71  operated  by  compressed  air,  and  376  elec- 
tric. The  anthracite  region  uses  mostly  the  steam-locomotive, 
while  the  electric  motor  is  more  common  in  the  bituminous 
district. 

All  traction-locomotives,  depending  solely  upon  the  adhesion 
of  their  drivers  to  the  rails,  are  limited  to  grades  of  4  per  cent. 
They  are  not  economical  on  steeper  grades,  except  for  short  dis- 
tances and  where  the  up-grade  is  approached  by  a  down-grade, 
enabling  the  train  to  acquire  momentum  enough  to  help  it  up  the 
steeper  grade.  - 

The  Haulage  Capacity  of  a  Locomotive. — This  is  measured 
by  the  weight  of  the  train  which  it  can  haul  upon  a  level  straight 
track.  Its  tractive  force,  or  its  draw-bar  pull,  is  the  tension  exer- 
cised by  the  locomotive  upon  its  first  connection  with  the  train. 


UNDERGROUND  HAULAGE  SYSTEMS.  271 

The  draw-bar  pull  must  exceed  the  total  frictional  resistance  of  the 
train  upon  the  track,  including  that  between  the  wheels  and  the 
rail,  the  resistance  of  the  curves  and  of  the  grade.  The  draw- 
bar pull  is  about  one  eighth  the  weight  of  the  locomotive.  Hence 
the  weight  of  a  locomotive  should  be  the  maximum  for  given 
minimum  dimensions. 

The  Tractive  Force. — The  tractive  force,  T,  of  a  steam-  or 
air-locomotive,  expressed  in  pounds,  is  measured  by  the  formula 
dT  =  i2k2sp,  wherein  d  is  the  diameter  of  the  driver  in  feet,  and 
k,  p,  and  s  have  values  as  in  Chapter  V  for  steam-engines.  The 
traction  of  a  locomotive,  expressed  in  pounds,  must  not  be  con- 
fused with  its  horse-power,  which  is  a  unit  of  dynamic  force, 
embracing  the  elements  of  weight,  distance,  and  time. 

The  train  resistances  R  are :  (i)  the  frictional,  which  in  mines  is 
not  less  than  30  Ibs.  per  ton  of  train,  and  equals  30  Y,  Y  being  the 
number  of  tons  weight  of  train  and  load ;  (2)  that  due  to  grade, 
which  is  2ogY,  g  being  in  feet  per  100;  (3)  that  due  to  the  curve, 
which  is  ^  Ib.  per  ton  per  foot  width  of  gauge,  z,  per  i°  curvature; 
and  4,  that  due  to  speed,  which  is  0.257+3,  V  being  the  speed  in 
miles  per  hour.  For  any  degree  of  curvature,  D,  and  weight,  w, 
of  cars  which  are  at  one  time  on  the  curve,  the  curve  resistance  is 
\wzD.  As  the  radius  of  a  i°  curve  is  5730  feet,  then,  when  the 

radius  of  the  curve,  p,   is  known,  the  curve  resistance  =7 16.2 — • 

Often  the  equivalent  grade  resistance  is  calculated;  that  is,  the 
percentage  of  grade  which  offers  the  same  resistance  as  the 
curve.  This  =  1.32!}. 

EXAMPLE. — A  locomotive  has  a  draw  bar  pull  of  2400  Ibs.  How  many 
tons  will  it  pull  up  a  grade  of  i  in  50  if  the  friction  is  32  Ibs.  per  ton? 
The  work  against  gravity  is  1 750,  or  45  Ibs.  per  ton,  making  a  total  resistance 
of  77  Ibs.  The  available  pull  being  2400,  the  gross  load  is  31  tons,  less  the 
weight  of  the  locomotive. 

The  influence  of  grade  in  reducing  the  haulage  capacity  of 
the  locomotive  is  shown  in  the  following  table.  For  any  other 
weight  of  locomotive  the  capacity  is  directly  proportional. 


272  MANUAL  OF  MlNIXG. 

HAULING  CAPACITIES  OF  IX>COMOTIVF.S  ON  VARIOUS  GRADES,  IN  TONS  OF  2000  LBS. 


Frictional 

Grades. 

Draw-bar 

Car  Resist- 

Weight. 

Pull  on 
Level. 

ance  per 
Ton  on 
LeveL 

Level. 

HPer 
Cent. 

i  Per 
Cent. 

iHPer 
Cent. 

2  Per 
Cent. 

2^Per 
Cent. 

( 

20 

23 

15 

1C 

8 

6-3 

5-2 

4,000 

500  ] 

3° 

IS 

II 

8.4 

6.7 

5-4 

4-5 

( 

40 

12 

9 

7.0 

5-7 

4-7 

4.0 

Frictional 

Grades. 

Draw  -bar 

Car  Resist- 

Weight. 

Pull  on 
Level. 

ance  per 
Ton  on 
Level. 

Level. 

Cent. 

3^  Per 
Cent. 

Cent. 

5  Per 
Cent. 

6  Per 
Cent. 

( 

20 

23 

4-2 

3-5 

3-0 

2.2 

1.6 

4,000 

5«M 

3° 

15 

3-8 

3-2 

2-7 

2.0 

1-5 

( 

40 

12 

3-4 

3-o 

2-5 

1.8 

1-4 

Comparison  of  Types  of  Locomotives. — At  first  the  steam- 
locomotive  was  used  underground,  but  the  objections  to  it,  which 
were  soon  recognized,  led  to  its  removal  from  gaseous  mines  or 
from  those  whose  ventilating  system  may  become  disturbed  by  the 
passage  of  the  locomotive  through  the  airway,  and  it  has  been 
supplanted  by  the  later  forms  of  motor-power. 

The  compressed-air  locomotive  is  the  only  cheap  form  of 
haulage  for  gaseous  mines.  It  introduces  no  element  of  dan- 
ger other  than  that  common  to  all  large  haulage  motors — the 
interference  with  intake  air-current  during  the  outward  passage 
of  the  locomotive. 

The  electric  locomotive  has  now  established  itself  in  coal- 
mines, and  may  be  regarded  as  a  safe  form  of  haulage  except 
in  special  cases.  It  may  be  had  of  three  different  types— the 
overhead-wire  trolley,  the  third-rail,  or  the  rack-and-pinion  when 
the  grade  is  excessive  for  the  simple  traction-engine. 

On  the  other  hand,  a  very  decided  objection  obtains  against 
the  air-  and  the  steam-locomotive,  and  to  a  less  degree  against  the 
electric  locomotive,  in  the  area  of  the  haulage-way  which  they 
occupy.  Unless  the  latter  is  excavated  to  a  large  area  the  passage 
of  the  locomotive  interferes  markedly  with  the  air-current.  When 


UNDERGROUND  HAULAGE  SYSTEMS. 


273 


- 


274  MANUAL  OF  MINING. 

travelling  with  the  current,  at  least  20  per  cent  more  air  enters 
the  mine  than  when  the  train  is  moving  out  against  it.  This 
may  be  averted  by  disconnecting  the  locomotive  haulage-way 
from  the  main  ventilation  system  of  the  mine  and  introducing 
the  inlet  current  at  a  point  beyond  the  inside  terminus  of  the 
locomotive  gangway.  The  cost  of  these  improvements,  how- 
ever, is  very  large. 

The  Steam-locomotive  has  not  as  ready  acceptance  in  under- 
ground work  now  that  it  had  in  former  days,  because  of  the 
objection  raised  to  the  presence  of  the  gaseous  products  which 
it  discharges  into  the  airway.  Whether  coal-burning  or  oil- 
burning,  there  is  given  off  a  considerable  percentage  of  car- 
bonic oxide  gas  as  well  as  smoke.  The  former  is  a  menace,  and 
the  latter  befouls  the  ventilation  besides  introducing  the  risk  of 
fire.  Moreover,  the  great  heat  and  moisture  in  the  exhaust 
tend  to  rot  the  timbers  and  to  soften  the  roof.  These  are  about 
the  same  objections  that  obtain  against  any  underground  steam- 
engine.  In  long  haulage-ways  steam-locomotives  can  be  used 
only  when  a  special  means  of  ventilation  is  provided.  There  are 
many  instances  where  lives  have  been  lost  by  the  inhalation  of 
the  gases  from  the  locomotive  and  underground  engines.  The 
steam-locomotive,  however,  is  less  expensive  than  the  air-engine 
and  very  much  lighter  in  weight. 

The  capacity  depends  on  the  driver  adhesion,  and  a  steam- 
locomotive  of  6"Xio"  cylinders  will  haul  28  tons  of  train  at  a 
rate  of  twenty  miles  per  day  up  a  grade  of  105  feet  per  mile  with 
600  Ibs.  of  coal.  One  of  10"  Xi4"  cylinders  has  a  duty  of  46 
tons  per  hour  at  28  miles  speed  on  grades  of  52  feet  per  mile. 
The  coal  consumption  is  1000  Ibs.  per  day. 

The  locomotives  are  made  of  a  shape  to  suit  the  mine  open- 
ing, for  narrow  gauge  (36  to  40  inches)  rarely  over-  78  inches 
high,  have  four  to  six  wheels  (for  curves  of  50  to  75  feet  radius), 
weigh  4  to  13  tons,  and  carry  125  to  350  gallons  of  water.  Their 
cylinders  are  from  5  Xio  to  10X14,  on  22-inch  to  28-inch  drivers 
running  over  16-  to  28-  Ib.  rails.  These  engines  cost  from  $2600  to 
$4000.  They  have  a  traction  of  from  150  to  600  tons  on  a  level. 


UNDERGROUND  HAULAGE  SYSTEMS.  275 

There  is  no  difference  in  the  price  between  the  wide-  and  the 
narrow-gauge  locomotive  of  the  same  design  and  size  of  cylinders. 
For  narrow  tunnels  locomotives  with  inside  cylinders  may  be  had. 
The  Compressed-air  Locomotive. — This  consists  of  one  or 
two  storage  tanks,  in  which  is  retained  compressed  air  sufficient 
in  amount  for  the  entire  run  of  the  trip,  and  two  cylinders  with 
the  usual  driver  connections.  When  the  air-pressure  is  as  high 


FIG.  112. — A  Mine  Locomotive. 

as  1000  Ibs.  per  square  inch,  the  high-pressure  tank,  placed 
along  the  side  of  a  larger  tank,  is  quite  small,  with  a  reducing- 
valve  between  them.  The  cylinder  receives  air  from  the  low- 
pressure  tank.  As  the  radius  of  action  depends  on  the  capacity 
of  the  tank,  the  latter  is  double  that  calculated  to  be  required  for 
the  trip.  The  cars  are  hauled  to  the  foot  of  the  slope,  or,  if  the 
grade  is  a  natural  one,  all  the  way  to  the  tipple.  The  locomotive 
goes  as  far  into  the  mine  as  the  height  of  the  seam  and  the  local 
conditions  will  admit. 

These  locomotives  are  operated  precisely  like  the  steam- 
locomotive,  but  are  of  greater  weight,  with  larger  bearings,  frames, 
cylinders,  wheels,  etc.,  though  the  general  pattern  is  much  the 
same.  The  early  objection  to  the  air-engines — the  fear  of  ex- 
plosion— is  no  longer  warranted.  Their  dimensions  are  about 
as  follows:  16  to  20  feet  in  length,  6  to  7  feet  in  width,  and 
5  feet  in  height.  The  cylinders  are  not  lagged  as  are  steam- 
cylinders;  indeed  their  exterior  surface  is  corrugated  to  increase 
the  exposure. 


MANUAL  OF  MINING. 


The  haulage  capacity  is  calculated  in  the  same  manner  as  for 
steam-locomotives,  the  percentage  sometimes  allowed  for  adhesion 
of  drivers  to  the  rails  being  one  fourth.  Their  tractive  efforts 
are  calculated  in  the  same  way;  after  ascertaining  the  mean 
effective  pressure  from  a  given  ratio  of  expansion  and  from  the 
size  of  the  cylinder,  the  volume  of  air  required  to  deliver  the  given 
power  at  a  known  rate  of  speed  is  found.  In  determining  the 
mean  effective  pressure,  the  second  table  in  Chapter  X  will  be 
found  convenient  for  conditions  in  which  the  air  is  at  a  pressure  of 
not  to  exceed  200  Ibs.  per  square  inch.  Above  that  the  formulae 
may  be  employed  for  the  purpose. 

The  following  table  shows  the  tractive  effort  of  certain  air- 
locomotives. 

TRACTIVE  EFFORTS  OF  COMPRF>SSED-AIR  LOCOMOTIVE. 


Cylinder. 

Diameter 

Weight 

Tractive  Effort  for  each  loo-lb.  Gauge  Cylin- 
der Pressure  at  Various  Cut-offs,  for 

Driver, 

Driver, 

Stroke, 

Inches. 

Inches. 

H  in., 

^in., 

Kin., 

Hin., 

?*in., 

y4in.. 

Hin., 

0.98 

0.95 

0.88 

0.80 

0.68 

0.51 

0.31 

5 

10 

24 

6,000 

1020 

990 

920 

835 

710 

530 

325 

6 

IO 

24 

8,500 

1470 

1425 

1320 

I2OO 

IC2O 

760 

445 

7 

12 

26 

13,000 

2200 

2150 

1990 

1810 

1540 

1140 

700 

8 

12 

26 

18,000 

2880 

2750 

2600 

2360 

2000 

1510 

900 

9 

14 

26 

25,000 

4340 

4140 

3840 

349° 

2960 

2220 

135° 

10 

14 

26 

32,000 

S28o 

5150 

4740 

431° 

3660 

2630 

1670 

ii 

16 

28 

42,000 

6770 

6450 

SQ«o 

5440 

4620 

347° 

2440 

12 

16 

28 

52,000 

8050 

7800 

7200 

6550 

5580 

4150 

2850 

Aside  from  its  low  efficiency,  the  main  objection  to  the  com- 
pressed-air locomotive  is  in  the  great  dimension  of  its  tanks. 
The  exhaust  is  laden  with  snow  if  the  ratio  of  expansion  has  been 
high,  for  the  air  is  usually  filtered  through  water  previous  to  com- 
pression, to  remove  the  dust.  This  charges  it  thoroughly  with 
moisture,  which  causes  the  trouble. 

Compound  pneumatic  locomotives  are  being  introduced  in 
order  to  utilize  the  expansive  effort  of  the  air.  They  have  an 
advantage  over  the  simple  locomotive  of  lower  piston  speed  for 


UNDERGROUND  HAULAGE  SYSTEMS.  277 

the  same  consumption  of  air.  A  higher  initial  pressure  is  then 
used. 

The -cost  of  a  compressed-air  plant,  with  motors,  boilers,  etc., 
is  $33,000  for  a  mine  having  an  output  of  900  tons  per  day  with 
a  mean  haul  of  2000  feet.  The  cost  of  haulage  per  ton  mined  is 
12.6  cents. 

The  storage  tanks  on  the  locomotive  are  charged  at  either  end 
of  the  line  from  stations  where  large  storage  tanks  are  located, 
which  latter  are  in  pipe  communication  with  the  air- compressor. 
The  capacity  of  these  stationary  tanks  exceeds  twice  the  capacity 
of  the  locomotives  comprising  the  system. 


FIG.  113. — A  Pneumatic  Locomotive. 

The  Electric  Locomotive. — This  is  the  most  advantageous 
form  of  electric  application  for  underground  work  both  in  point 
of  economy  and  convenience.  Its  introduction  has  been  of  slow 
growth  in  coal-mines  because  of  the  conservatism  enforced  upon 
engineers  by  adverse  legislation,  which  has  always  compelled 
them  to  go  slow  in  the  introduction  of  any  new  form  of  machinery. 
Since  its  first  installation  it  has  established  itself  and  is  more 
popular  than  the  compressed-air  motor,  because  it  is  more  com- 
pact and  more  economical.  Its  compactness  enables  it  to  serve  in 
entries  too  small  for  compressed  air  by  engines  and  too  sinuous 
for  cable-ways.  The  results  from  its  introduction  have  not  proved 
any  more  destructive  or  injurious  to  life  than  from  other  forms  of 
haulage  power,  and  it  is  used  in  any  gangway  where  the  naked 
lamp  can  be  carried.  It  possesses  the  same  advantage  as  other 
locomotives  in  having  a  practically  unlimited  radius  of  action,  and 
has  an  additional  advantage  in  being  cheaper  to  operate  than  its 


278  MAXUAL  OF  MINING. 

competitive  traction  machinery,  having  nothing  to  get  out  of 
order.  Because  of  this  little  trouble  ensues. 

Two  systems  of  electric  locomotives  are  used;  one  with  over- 
head-trofley  wire  and  the  other  with  the  third  rail. 

The  mechanism  of  the  underground  locomotive  is  similar  in 
kind  to  that  of  the  familiar  street-car  trolley  in  which  hard-drawn 
copper  wire  is  laid  along  the  roof  of  the  road.  This  form  of 
the  overhead  wire  on  which  the  trolley  of  the  locomotive  arm 
rests  conducts  the  current  to  the  motor,  whence  it  may  return  by 
the  rail  to  the  generator. 

The  third- rail  locomotive  has  a  rack  between  the  two  main 
rails  in  which  engage  two  sprocket-wheels  driven  by  the  elec- 
tric motor  by  means  of  suitable  gearing.  These  sprocket- 
wheels  serve  the  double  purpose  of  driving  the  locomotive 
along  the  track  and  taking  the  electric  current  from  the  third 
rail  to  feed  the  electric  motor.  These  machines  are  of  recent 
introduction. 

Locomotive  Details. — It  is  impossible  to  indicate  the  average 
dimensions  of  the  locomotive,  because  each  one  represents  a 
special  type  designed  for  local  conditions  only.  Any  desired 
height,  power,  or  gauge  may  be 'procured.  Its  height  is  rarely 
more  than  4  inches  above  the  tops  of  the  wheels;  its  length  is  n 
feet,  and  width  57  inches,  for  a  gauge  of  36  inches.  Its  average 
weight  is  10  tons.  This  would  have  50  horse-power  capacity 
and  would  be  capable  of  taking  25  cars  of  4.5  tons  gross  weight 
each  on  the  level. 

The  driving  motors  are  made  as  small  as  is  consistent  with 
good  design,  and  are  suspended  as  low  as  safety  warrants  in  the 
space  between  the  wheels.  They  combine  both  high  power  and 
good  speed;  their  mechanism  is  of  simple  character  and  easily 
accessible.  A  housing  affords  protection  from  dust  or  injury. 
They  are  enclosed  in  a  heavy  cast-iron  frame  firmly  bolted 
and  carrying  axle-boxes.  To  each  axle  a  steel-clad  motor  is 
geared.  The  gears  are  of  the  best  cast  steel  and  designed  for 
maximum  strains  which  are  sure  to  occur  when,  in  the  darkness 
of  the  mines,  it  may  become  necessary  to  reverse  from  full  speed. 


UNDERGROUND  HAULAGE  SYSTEMS.  279 

Sand  is  carried  in  boxes,  fore  and  aft,  fitted  with  spouts  in  front 
of  each  wheel;  the  brake  is  designed  to  be  applied  by  one  hand. 
The  screw-brake  is  preferred  to  gear  or  ratchet. 

The  wheels  are  generally  of  chilled  iron,  30  inches  in  diame- 
ter, and  the  trucks  of  the  prevailing  gauge.  Steel  tires  are  more 
expensive,  but  furnish  a  greater  tractive  effort  than  chilled 
wheels.  The  length  of  wheel-base  is  governed  by  the  existing 
curves.  This  is  5  feet  for  a  gauge  of  36  inches  and  a  radius  of 
25  feet. 

The  trolley-pole  is  reversible  and  adjustable  to  maximum  as 
well  as  minimum  heights.  It  may  be  placed  on  either  side  or 
end  of  the  car  as  conditions  require,  the  centre  of  the  socket 
being  over  the  gauge  line  and  on  the  ditch  side,  to  reduce  the 
element  of  danger.  The  motor  is  of  the  slow-speed  four-pole 
type  with  windings  suited  to  the  given  line  pressure  and  a  speed 
of  eight  miles  per  hour,  taking  into  consideration  the  line  loss. 
If  it  happens  that  the  line  loss  is  less  than  was  assumed,  the  loco- 
motive speed  will  exceed  this. 

The  Rheostat. — The  controlling  device  is  of  the  five-point 
rheostatic  type  of  electric  railway  motor,  so  designed  as  not  to 
heat  to  the  danger-point  when  using  the  maximum  current.  This 
form  of  diverter,  or  rheostat,  is  necessary  because  of  the  great 
variation  in  load  and  grade  which  the  locomotive  encounters  in 
a  run. 

The  resistance-box  is  placed  in  front,  where  it  will  receive  a 
strong  circulating  current  of  air.  Powerful  magnetic  blowouts 
are  always  provided  to  insure  against  sparking. 

The  Draw-bar  Pull  of  an  Electric  Locomotive.  —  This  is 
about  50  Ibs.  per  horse-power  of  the  motor,  and  the  weight  of  the 
locomotive  at  the  ordinary  speed  of  running  is  about  400  Ibs.  for 
each  horse-power  of  motor,  making  the  tractive  effort  of  an  elec- 
tric locomotive  about  one  eighth  of  its  weight.  For  continuous 
work  the  effective  tractive  effort  should  be  assumed  as  20  per 
cent  less  than  this  rated  capacity.  This  will  prevent  overheating 
of  the  wires. 

The  average  bar  pull,  P,  of  different  weights,  W,  of  loco< 


280 


MANUAL  OF  MINING. 


motives  expressed  in  pounds  on  varying  grades,  g,  is  expressed 
as  follows: 

P  =  W  (0.125  —g)  =  horse- power  of  locomotive X (50— 4^).  A 
ten- ton  locomotive  can  pull  1900  Ibs.  on  a  3  per  cent  grade  or 
1300  Ibs.  on  a  6  per  cent  grade,  as  against  2500  Ibs.  on  a  dead 
level.  The  starting  draw-bar  pull  is  nearly  60  per  cent  greater 
than  the  ordinary  running  rate  when  sand  is  used. 

Locomotive  Rating. — The  length  of  time  for  which  a  motor 
can  give  a  stated  ampere  rating,  7,  is  based  on  the  expression  tP> 


II- 


ELECTRIC  MINING  LOCOMOTIVE  CURVE 
BOOVOITS 


\ 


FlG.  114. — The  Power  Curves  of  a  Locomotive. 


in  which  t  is  the  number  of  minutes  of  running,  and  7  the  number 
of  amperes  for  regular  work.  Every  overload  causes  heating 
beyond  the  allowed  limit.  This  varies  with  the  square  of  the 
current.  The  rating  of  a  given  locomotive  being  known,  ft  is 
possible  thus  to  ascertain  its  hauling  capacity. 


UNDERGROUND  HAULAGE  SYSTEMS.  281 

Thus  a  motor  rated  at  50  amperes  for  18  minutes  can  work 
only  4|  minutes  at  100  amperes.  The  length  of  haul  also  has 
an  influence  on  the  capacity  of  the  motor.  One  intended  to  pull 
a  maximum  load  for  a  long  period  must  have  a  greater  capacity 
than  one  which  is  operating  for  short  hauls,  allowing  time  to 
cool  between  stops. 

Notation. — A  system  has  been  adopted  for  notation  of  elec- 
tric locomotives  similar  to  the  Whyte  system  for  steam-locomo- 
tives. The  number  of  wheels  under  the  locomotive,  the  number 
of  motors  and  horse-power  of  each,  and  the  number  of  pairs  of 
drivers  geared  to  the  motor  are  indicated  in  the  order  named. 
The  letter  indicates  the  number  of  drivers:  B,  for  one-geared 
axle;  C,  two  pairs  of  drivers;  D,  three  axles  geared  to  motors. 
4/o-C  means  4  wheels,  2  motors  of  50  H.P.  each,  and  2  geared 
axles. 

Storage-battery  Locomotives. — Owing  to  the  numerous  acci- 
dents from  contact  with  the  wires,  whereby  the  mule  becomes  an 
unnecessary  evil  in  the  operation  of  the  mine,  the  animal  may  be 
eliminated  from  the  haulage  systems  by  employing  manual 
labor  for  pushing  the  car  from  the  parting  into  the  entry.  A  com- 
bined storage  and  trolley  locomotive  has  also  been  devised  for 
the  purpose  and  employed  with  success.  Storage  batteries  obvi- 
ate the  necessity  for  wires,  but  as  yet  are  too  expensive  and  also 
too  heavy  for  efficient  use.  A  25-lb.  battery  will  furnish  one 
horse-power  hour;  and  a  space  of  40  sq.  ft.  will  accommodate  250 
elements,  which  would  be  able  to  give  n  horse-power  during  a 
full  ten-hour  shift.  This  weight  and  the  adhesion  is  more  than 
twice  that  of  the  steam-locomotive  of  equal  horse-power,  but 
requires  a  heavier  rail. 

EXAMPLE. — It  is  desired  to  move  sixty  cars  of  a  capacity  of  3  tons  each  and 
weighing  i  ton  over  a  track  having  the  following  profile,  favoring  the  loaded 
cars:  1920  feet  of  1.3  per  cent  grade,  1000  feet  of  1.8  per  cent  grade,  600  feet 
of  2  per  cent  grade,  400  feet  of  2.5  per  cent  grade,  425  feet  of  3.5  per  cent 
grade,  and  1000  feet  of  0.9  per  cent  grade.  The  locomotive  rating  is  64, 
voltage  500,  amperes  32  for  each  motor,  and  speed  6  miles  an  hour. 

The  maximum  resistance  is  on  the  steepest  grade,  or  4^  per  cent. 

If  the  car  friction  be  assumed  at  i  per  cent,  the  cars  will  require  5400  Ibs 


282  MANUAL  OF  MINING. 

draw-bar  pull  to  do  this  work.  To  provide  for  the  friction  of  the  locomotive 
itself,  which  is  1280  Ibs.,  the  total  tractive  effort  must  be  6680  Ibs.  With 
cast-iron  chilled  wheels  the  weight  of  the  locomotive  must  be  53,280  Ibs. 
Dividing  the  draw-bar  pull  by  50  shows  that  the  locomotive  must  be  equipped 
with  io8-horse-power  motors  to  do  the  work  without  overload.  For  con- 
tinuous work  we  have,  for  the  two  motors, 

P 

-  =  2X32. 

T=the  total  time  that  should  be  taken  for  each  trip.  Calculating  for  the 
several  sections  the  time  of  haul  in  minutes  and  the  required  amperes, 
respectively,  as  below,  with  a  voltage  of  500: 

Time       =      3.8             2            i.i           0.8           0.8  2  minutes 

Amperes  =     114           128          134          148           182  100 

and  tP          =  41,400      32,768      19,751      17.523      26,500  20,000 
we  have              //2=  165,942;    (64)^=165,942;  and  ^=40. 

Cost  of  Electric  Haulage. — The  life  of  the  average  electric- 
haulage  plant  is  estimated  as  twenty  years,  and  hence  in  dis- 
tributing the  cost  of  haulage  an  allowance  of  5  per  cent  depreci- 
ation per  year  divided  by  the  number  of  working  days  per  year 
will  give  the  allowance  to  be  made  for  this  item  per  day. 

Comparing  some  instances  of  the  cost  of  electric  haulage  with 
that  of  mules,  the  evidence  preponderates  in  favor  of  the  former. 
On  the  item  of  track  cost  alone  it  shows  a  saving  of  25  per  cent 
on  the  investment  per  year. 

Self-acting  Planes. — Conditions  occasionally  occur  when 
gravitation  can  be  utilized  to  deliver  the  loaded  car  to  the  bottom 
of  the  plane.  A  double  line  of  rails  is  laid  the  entire  length  of 
the  track  for  two  trains  of  cars.  A  rope  connects  them  and 
passes  around  the  sheave  at  the  top  of  the  incline  fitted  in  a  recess 
above  the  landing  where  the  cars  are  started.  In  some  cases  a 
single  rope  is  used,  the  sheave  being  horizontal  and  its  axis  ver- 
tical; in  other  cases  a  drum,  whose  axis  is  horizontal,  carries  two 
ropes  wound  in  opposite  directions,  each  one  having  one  or  more 
cars  attached  to  it  when  desired.  A  single  rope  may  be  used  ori 
the  drum  with  five  or  six  extra  coils  upon  the  drum  between 
the  two  branches  of  rope  to  give  friction  enough  to  prevent  the 
rope  from  slipping.  This  form  of  plane  is  used  in  haulage- ways, 


UNDERGROUND   HAULAGE  SYSTEMS.  283 

and  occasionally  in  individual  rooms  for  the  delivery  of  one  or 
more  cars  at  a  time. 

Sometimes  three  rails  are  laid  with  a  suitable  turnout  mid- 
way along  the  incline.  This,  however,  is  false  economy,  as  com- 
pared with  the  full  four- rail  line. 

The  lowest  gradient  at  which  the  plane  is  self-acting  depends  on 
the  state  of  the  road  and  the  load,  but  under  the  most  favorable 
conditions  the  gradient  requires  to  be  about  1.25  inches  per  yard, 
or  i  in  29.  At  about  6  per  cent  pitch  it  works  freely.  The  most 
satisfactory  grade  is  i  in  6.  When  the  grade  is  as  large  as  i  in  3 
there  is  an  excess  of  motor  force  above  the  resistance,  and  a  suitable 
brake  must  be  supplied  to  control  the  speed.  Some  inclines  have 
a  counterbalance-car.  The  average  safe  velocity  on  such  a  plane 
is  400  feet  per  minute.  A  two-car  plane  on  10°  incline  requires  a 
f -inch  rope.  On  a  45°  slope  the  rope  is  f  inch  ihxdiameter  for 
three  cars. 

The  lower  portion  of  the  inclination  is  flattened  somewhat  to 
diminish  the  momentum  of  the  descending  cars,  and  in  some 
cases  a  reverse  grade  is  built  in  order  to  check  the  train.  The 
speed  of  the  car  or  train  is  regulated  by  a  brake  applied  on  the 
band  bolted  over  the  sheave- wheel.  It.  is  of  ihc  usual  type,  fitted 
with  good  durable  surface  having  a  high  coefficient  of  friction. 
The  safety  attachments  provided  on  cars  are  essentially  con- 
fined to  some  form  of  trailing-block. 

Single-track  self-acting  inclines  for  single  cars  have  a  trailing 
counterbalance-car  to  regulate  the  speed,  with  a  turnout  at  the 
middle  of  the  line. 

Engine  Planes. — When  the  slope  is  against  the  load  and  a 
locomotive  cannot  be  employed  to  advantage,  engine  planes  are 
used.  A  stationary  engine  with  drum  is  located  at  the  head  of 
the  plane  to  raise  the  cars  in  the  train  at  any  speed  which  the 
timbering  of  the  roadway  will  permit.  If  the  traffic  is  large 
enough  to  warrant  it,  the  engine  is  in  continual  motion,  the  drum 
being  thrown  in  or  out  of  gear  when  desired.  If  the  drum  may 
turn  freely,  it  will  pay  out  the  rope  for  the  descending  cars,  and  is 
geared  to  pull  them  up.  On  a  grade  of  1.7  per  100,  gravity  will 


284 


MANUAL  OF  MINING. 


UNDERGROUND  HAULAGE  SYSTEMS.  285- 

take  the  loaded  cars  down  with  a  reasonable  velocity  (empties  on 
a  2.25  grade),  pulling  the  rope  behind  them.  On  a  10  per  cent 
grade  a  brake  will  be  necessary  for  the  empty  cars. 

The  track  may  be  single  or  double,  according  to  the  output 
desired.  On  single-track  planes  the  engine  is  non-reversing.. 
The  cars  usually  travel  in  trains,  ten  to  thirty  cars  to  a  train,  each 
in  charge  of  a  conductor,  who  operates  a  dead-fall  timber  block  to 
hold  the  train  while  the  cars  are  being  shunted.  The  size  of 
the  engine  can  be  calculated  as  for  hoisting,  by  determining  the 
amount  of  power  requisite  to  overcome  the  vertical  component 
of  the  weight  of  the  rope  and  train  and  the  horizontal  resistance 
of  the  friction.  The  engine  plane  is  eminently  suited  to  an  oil- 
engine or  an  electric  installation.  It  is  well  adapted  for  the 
delivery  from  side  entries  at  different  levels,  and  may  be  used  on 
slight  curves  by  curving  the  rope  on  iron  guide-wheels.  Ordi-. 
narily  the  rope  will  last  four  years.  A  14"  X3o"  engine  with  a 
3-ton  fly-wheel,  T\-inch  steel  rope  on  a  plane  4600  feet  long  and 
80  feet  rise,  has  a  daily  output  of  950  tons  of  mineral  in  trains  of 
25  to  30  cars. 

Rope-haulage  Systems. — The  two  classes  of  rope-haulage 
systems  are  known  as  the  tail-rope  and  the  endless  rope.  Both 
are  extensively  used  in  horizontal  seams,  and  while  each  has  its 
special  adaptation  to  a  given  grade,  local  conditions  may  make 
it  possible  of  application  under  steeper  grades.  The  tail-rope 
can  be  employed  for  grades  not  exceeding  3  in  100,  either  with 
or  against  the  loaded  cars. 

The  Tail-rope  System. — This  consists  of  a  stationary  engine 
with  two  drums  which  are  thrown  alternately  in  and  out  of  gear. 
From  each  drum  is  carried  a  rope.  The  main-rope  has  a  length 
equal  to  that  of  the  road,  and  is  fastened  to  the  front  end  of  the 
train.  The  tail-rope,  of  double  this  length,  passes  from  the  drum 
around  a  sheave  at  the  end  of  the  road  and  up  to  the  rear  end  of 
the  train. 

When  hauling  the  loaded  train  out,  the  main- rope  drum  is 
thrown  into  gear  and  that  carrying  the  tail-rope  is  free.  The 
loaded  cars  are  then  driven  to  the  outlet  or  the  point  of  distribu- 


286 


MAX  UAL   OF   MIXIXG. 


UNDERGROUND  HAULAGE  SYSTEMS.  287 

tion,  dragging  behind  them  the  tail-rope.  When  the  train  has 
been  replaced  by  empty  cars,  the  main- drum  is  released  and  the 
tail-drum  engaged.  The  engine  is  started  and  the  cars  are  de- 
livered into  the  mine.  The  lower  sheave  is  a  grip-pulley  similar 
to  that  used  in  the  power  transmission,  Fig.  118.  The  main- 
rope  is  rarely  over  f  inch  in  diameter.  The  tail-rope  is  usually 
an  old  discarded  main-rope.  Swivel  connections  are  used  from 
the  rope  to  the  train. 

The  roadway  is  usually  a  single  track,  and  the  sheave  is  ver- 
tical with  the  tail-rope  carried  along  the  roof,  while  the  main- 
rope  is  supported  along  the  floor.  The  length  of  haul  is  limited 
only  by  the  engine  power  and  the  track  resistances.  The  most 
advantageous  pitch  is  about  3  per  cent  outward  from  the  workings. 
The  velocity  of  haul  may  reach  as  high  as  10  miles  an  hour. 
Each  trip  may  take  a  train  of  from  10  to  50  cars,  usually  ac- 
companied by  one  trailer  in  charge  of  a  conductor. 

This  system  is  very  much  used  in  American  collieries.  It 
is  the  best  plan  for  operating  branch  ways.  Each  branch  has 
its  own  rope  passing  over  sheaves  at  the  ends.  When  a  train  is 
to  be  hauled  out  of  the  room  to  the  point  of  discharge,  the  principal 
ropes  are  opened  at  the  proper  points  and  connections  with  the 
branch  rcpes  are  made;  meanwhile  the  other  branches  are  idle. 
This  is  an  inexpensive  plant  to  build  and  to  maintain.  It  is 
advantageous  for  undulating  roadways  and  has  very  largely 
replaced  animal  power. 

An  engine  of  i8"X3o"  cylinder  with  75  Ibs.  steam-pressure 
draws  thirty  trains  of  17  cars  each  in  a  day  over  a  slope  of  2800 
feet  long  with  a  grade  of  i  in  200.  The  drums  are  4.5  feet  in 
diameter,  the  main-rope  is  f  inch  and  ihe  tail  rope  \  inch  diame- 
ter. The  velocity  of  travel  varies  between  8  and  n  feet  per 
second. 

A  tail-rope  p'ant  for  1000  tons  daily  for  a  mile  of  haulage 
will  require  about  130-!!.?.  boiler  and  engines;  6000  feet  of 
main-rope  i  inch  diameter;  16,000  feet  of  f -inch  tail-rope;  600 
rollers;  500  cast-iron  sheaves;  10  return  sheaves  and  drums. 
The  first  cost  would  be  nearly  $7500,  and  the  maintenance 
about  $20  per  day. 


288  MANUAL  OF  MINIXG. 

The  Endless-rope  System. — This  system  is  much  in  vogue. 
A  rope  or  a  chain  receives  continuous  motion  in  one  direction 
from  a  wheel  or  drum,  and  draws  the  cars  out  on  one  line  and  in 
on  the  other.  The  tension  requisite  to  carry  the  load  is  arti- 
ficially produced  by  a  carriage  provided  with  some  form  of  coun- 
terpoise. It  requires  a  double-track  line  for  communication 
and  considerably  less  rope  than  in  the  tail-rope  system.  Cars 
can  be  attached  at  any  point  on  the  line,  but  branch  connections 
are  not  possible  with  this  system. 

The  wire  rope  or  chain  cable  may  rest  on  the  top  of  the  car 
or  be  supported  below  it.  In  the  cheapest  and  most  universal 
method  the  chain  rests  in  forks  riveted  on  the  top  of  each  car. 
For  uniform  grades  and  sharp  curves  the  endless  rope  may  be 
suspended  above  the  cars,  to  which  it  is  attached  by  short  lengths 
of  chain.  If  the  chain  cable  is  underneath  the  cars,  it  runs  on 
rollers,  connections  being  made  by  hooking  a  short  length  of 
chain  from  each  car.  With  a  rope  running  over  rollers  or  sheaves 
on  the  floor,  like  the  surface  cable  roads,  connection  is  made  by 
Si  hand-grip  attached  to  the  car. 

The  engine  of  the  double-track  system  is  in  continual  rotation, 
^driving  a  shaft  fitted  with  a  heavy  fly-wheel,  and  with  a  friction- 
clutch  for  emergencies.  From  the  thaft  a  belt  or  rope  drives  a 
.parallel  shaft  with  the  sheave  about  which  the  endless  rope 
travels.  A  similar  sheave  at  the  far  end  of  the  line,  with  a 
balance-car  or  tightener,  constitutes  the  entire  mechanism.  The 
speed  of  the  rope  rarely  exceeds  four  miles  per  hour. 

The  cars  are  attached  singly  at  the  landings  as  fast  as  they 
arrive.  The  connection  is  made  by  wrapping  a  few  turns  of 
chain  about  the  rope  and  hooking  the  ends  on  by  the  hand-grip. 
The  latter  is  more  convenient  and  secure  than  the  former,  par- 
ticularly if  the  cars  are  in  irain.  The  grip  is  a  pair  of  clamps 
attached  to  one  end  of  the  car,  the  end  of  the  movable-jaw  lever 
(being  closed  on  the  rope  by  a  ring  slipping  on  to  its  mate.  Rais- 
ing the  ring  opens  the  clamp  and  drops  the  rope.  The  car  is 
released  and  switched  to  its  proper  track.  In  the  same  way  it 
•can  be  as  easily  clamped  on  the  rope  to  pull  the  car.  For  a 
train  of  cars  a  leading  grip-car  is  attached  by  a  gripman,  who 


UNDERGROUND  HAULAGE  SYSTEMS.  289 

takes  the  train  out  or  in.  Clevis  couplings  are  also  used  for 
single  cars  and  at  the  front  and  rear  ends  of  trains.  Sometimes 
a  "knock-off"  hook  is  used  at  the  rear  to  aid  in  uncoupling  while 
strain  is  on  the  rope. 

Increasing  the  Tension  of  the  Drums  at  the  Driving  End. — 
The  difference  in  tensions  due  to  the  two  loads  pulling  upon  the 
two  branches  of  rope  is  slight.  The  amount  of  friction  at  the 
sheave  must  at  least  equal  the  difference  of  loads.  Should  any 
excessive  loading  or  resistance  occur,  the  sheave  will  slip  and  fail 
to  drive.  A  tighter  grip  must  be  taken  on  the  rope  by  the  sheave 
to  communicate  power.  It  would  also  prevent  excessive  wear 
of  the  rope.  This  is  possible  where  the  sheave  has  several  grooves. 
The  number  of  coils  will  be  such  as  will  provide  the  necessary 
friction  to  prevent  the  rope  from  slipping  off.  This  may  be 
determined  by  the  formulae  for  power  transmission.  Two  drums 
in  tandem  Serve  much  better,  each  one  having  four  or  five 
grooves,  the  rope  making  a  half-coil  alternately  upon  each  drum 
in  each  groove  before  returning  along  the  other  trackway. 
They  are  both  geared  to  the  engine  for  equal  speed. 

The  Balance-car. —  The  variations  of  load  and  changes  of 
temperature  produce  varying  tensions  and  elongations  of  the 
rope.  To  insure  a  uniform  tension  the  sheave  at  the  other  end 
of  the  line  is  mounted  on  a  carriage  and  capable  of  sliding  freely. 
Beyond  it  is  a  sunken  pulley  over  which  is  the  rope  from  the  car- 
riage to  a  suspended  counterpoise.  When  the  loads  on  the  rope 
become  heavy,  the  carriage  is  pulled  by  the  pendent  weight  to 
increase  the  rope  tension.  With  a  light  load  the  weight  of 
the  rope  draws  the  carriage.  If  the  grades  favor  the  load,  the 
carriage  is  at  the  end  of  the  run  of  the  loaded  cars,  the  power 
being  inside  of  the  mine  at  the  top  of  the  empty  track.  On 
aerial  tramways  the  same  regulating  device  is  used,  and  in 
Fig.  118  is  an  illustration  of  one. 

Instead  of  the  plain  grooved  sheave  there  may  be  a  grip- 
pulley,  several  feet  in  diameter,  carrying  radial  forked  arms  to 
seize  the  rope.  The  arms  are  capable  of  being  lengthened  or 
shortened  as  the  rope  stretches  and  drags.  When  the  arms  have 


290  MANUAL   OF  MINING. 

been  extended  to  their  limit,  a  short  length  of  rope  or  a  few  links 
of  cable  are  removed  and  the  forks  readjusted. 

An  i8"X42<;  Corliss  engine,  with  two  36-inch  driving-pulleys,. 
1 8-inch  belts  to  a  pair  of  84- inch  pulleys,  and  two  pinions  of 
18  inches  diameter  with  26  teeth,  gearing  into  72-inch  spur-wheels 
with  91  teeth,  drives  two  72-inch  rope-drums.  The  drums  with 
four  grooves  each  operate  an  endless-rope  system  one  and  a 
half  miles  long,  delivering  1200  tons  daily  over  an  average  road. 
The  fuel,  labor,  etc.,  of  a  45<>ton  endless  ropeway  is  about  $7.50 
per  day.  The  initial  cost  is  about  $6700  for  a  plant  of  4200  feet 
double  track  (1500  feet  having  a  rise  of  40  feet,  the  rest  being 
level).  The  rope  is  |  inch  diameter. 

If  only  a  single  track  is  possible  in  the  haulage- way,  the  oper- 
ation is  intermittent.  The  engine  alternately  pulls  the  rope 
with  its  train  inward  and  outward,  being  reversed  for  each  trip. 
This  resembles  the  tail-rope  system  with  the  two  ropes  linked 
together.  One  line  of  rope  is  supported  on  overhead  sheaves, 
and  the  other,  to  which  the  cars  are  attached,  rests  on  the  floor 
rollers.  The  cars  are  in  train. 

The  road  may  be  more  readily  extended  with  this  system 
than  with  the  tail-rope.  It  costs  less  for  power  and  can  be  carried 
around  sharper  curves  and  over  steeper  grades.  It  can  be  used 
with  practically  every  condition  of  roadway  with  greater  econ- 
omy than  any  of  the  other  systems.  The  delivery  of  cars  is 
uniform  and  continuous,  regardless  of  the  length  of  haulage. 
The  amount  of  car  and  rope  repairs  is  less  than  in  the  tail-rope 
system.  The  risk  of  accident  to  men  along  the  roadways  or 
damage  to  property  is  also  less,  because  of  the  slow  speed. 
Though  applicable  to  sinuous  undulating  roads,  it  is  best  adapted 
to  uniform  grades  throughout  the  entire  length  of  the  roadway, 
because  of  the  varying  load  which  would  otherwise  be  thrust 
upon  the  engine  where  the  grades  are  steep  and  variable.  This 
would  be  particularly  true  when  the  intermittent  attaching  of 
cars  is  very  irregular. 

„   Supporting   the    Haulage-rope. — Rollers    and  sheaves    along 
the  haulage- way  support  the  rope  to  reduce  its  wear  and  to  dimin- 


UNDERGROUND  HAULAGE  SYSTEMS.  291 

ish  the  frictional  resistance.  They  are  placed  midway  between 
the  rails  on  suitable  blocks  spiked  to  the  ties,  and  20  to  30  feet 
apart.  They  are  also  employed  for  supporting  the  tail-rope 
near  the  roof  by  being  suspended  from  the  roof-timbers.  The 
rollers  are  usually  of  wood,  12  to  18  inches  long,  5  inches  in 
diameter,  and  carefully  turned  with  an  iron  journal  for  support. 
They  are  spaced  on  a  level  track  from  15  to  20  feet  apart  (closer 
together  if  the  rope  is  large  or  the  grade  steep).  The  bearings 
also  are  of  oak,  well  oiled.  Hollow  rollers  may  be  obtained  of 
cast  iron  with  curved  faces,  to  keep  the  rope  as  central  as  possible. 
They  are  shorter  than  the  wooden  rollers  and  offer  much  less 
resistance. 

Deflecting  the  Main-rope  Around  the  Curves. — This  is  accom- 
plished by  a  number  of  small  sheaves  laid  on  the  centre  line  of 
the  track,  concentric  with  the  rails.  They  are  of  a  diameter  of  10 
to  15  inches  and  are  spaced  3  feet  or  so  apart  to  give  a  slight 
deflection  of  each  one  for  the  entire  length  of  the  curve.  On 
reverse  curves  they  are  laid  nearer  the  inside  rail  than  the  outer 
rail,  with  their  axis  somewhat  inclined  and  leading  outward. 
The  formulae  for  determining  their  spacing  are  to  be  found  on 
page  177. 

Short  vertical  rollers  are  also  placed  on  the  ties  when  the  rope 
is  to  be  deflected  around  a  gentle  curve.  For  a  sharp  curve  the 
rollers  are  frequently  placed  at  the  side  of  the  track,  with  their  axes 
slightly  inclined  from  the  vertical  to  carry  the  rope.  These  are 
spaced  3  feet  apart  outside  of  the  inner  rail  of  the  curve.  They 
are  usually  of  cast  iron,  chilled,  with  an  upper  flange.  Guide- 
blocks  are  laid  at  intervals  from  the  inner  rail  toward  the  sheave 
on  all  simple  curves  in  order  to  guide  the  rope  into  its  position  in 
the  groove,  and  also  to  prevent  its  catching  in  the  rail  as  the 
train  moves  around  the  curve.  These  blocks  are  of  oak,  spiked 
to  the  ties. 

Supporting  the  Tail-rope. — Tail-rope  sheaves  are  hung  on 
the  posts  at  the  sides  of  the  haulage- way,  25  to  30  feet  apart. 
For  these  a  perfect  alignment  is  necessary,  in  order  to  reduce 
friction  and  increase  the  security  to  the  rope  in  the  groove.  They 


2Q 2  MANUAL  OF  MINING. 

are  of  chilled  cast  iron  or  case-hardened,  having  wooden  filling 
in  the  grooves,  and  are  from  10  to  16  inches  in  diameter. 

The  tail-rope  is  turned  out  of  a  side  heading  by  a  horizontal 
guide-sheave  of  a  diameter  as  large  as  convenient.  Its  pattern 
is  like  that  of  the  clip-pulley  described  under  Power  Transmission, 
Chapter  IX.  A  large  diameter  is  necessary  because  the  angle  of 
contact  exceeds  90°. 

An  electric  system  along  the  tramways  is  requisite  for  safety 
as  well  as  for  signalling,  though  the  malicious  destruction  of 
insulation,  etc.,  has  caused  its  abandonment  by  many  operators. 

EXAMPLES. — i.  A  one-mile  endless  rope,  travelling  at  two  miles  an  hour 
over  an  average  grade  of  3  in  100,  delivers  50  tons  per  hour.  Required  the 
total  resistance  on  the  line  and  the  size  of  the  cylinders  under  50  Ibs.  boiler 
pressure  and  160  feet  piston  speed.  Each  car  weighs  800  Ibs.  and  carries  2000 
Ibs.  Along  the  line  are  distributed  25  loaded  and  25  empty  cars,  about 
100  feet  apart,  and  a  ton  is  delivered  every  36  seconds;  with  a  coefficient  of 
friction  of  0.02,  the  frictional  resistances  on  the  halves  of  the  line  are  1400 
and  400  Ibs.  respectively.  To  raise  25  tons  up  the  plane  requires  a  force 
of  3  per  cent  of  50,000,  or  1500  Ibs.;  the  gravity  component  of  the  cars  is 
600  Ibs.  The  gravity  components  of  the  cars  and  the  rope,  up  and  down, 
balance  each  other,  leaving  the  work  of  the  engine  to  be  that  of  overcom- 
ing 1500+  1800  (the  drag  of  the  rope  is  assumed  at  3000),  which,  carried  at  a 
rate  of  176  feet  per  minute,  requires  1,109,800  ft. -Ibs.,  or  33.6  I.H.P.  The 
piston  diameter,  k,  is  7$  inches,  and  for  130  strokes,  s,  nearly  15  inches. 
If  the  driving-sheave  is  6  feet  diameter,  it  makes  9.3  revolutions  per  minute, 
and  is  therefore  geared  i  to  7. 

2.  A  haulage-engine  having  cylinders  24"X48",  at  the  head  of  a  plane 
of  10  per  cent  grade,  is  directly  connected  to  its  drum.  With  an  effective 
steam  pressure  of  45  Ibs.  per  square  inch,  a  piston  speed  of  350  feet  per 
minute,  and  a  train  speed  of  8  miles  an  hour,  what  is  the  size  of  the  drum 
and  the  capacity  of  the  plant? 

The  total  piston  pressure  is  27,150  Ibs.,  assuming  a  modulus  of  §,  which, 
at  350  feet  per  minute  and  a  train  speed  of  704  feet  per  minute,  represents 
a  continuous  load  of  13,497  Ibs.  This  tension  upon  the  hauling-rope,  due 
to  the  component,  parallel  to  the  plane,  of  the  weight  of  the  rope,  cars,  and 
load,  plus  their  friction,  is  therefore  limited  to  13,497  Ibs.  For  this  working 
load  the  rope  may  be  ij  inches  diameter,  weighing  2  Ibs.  per  running  foot; 
the  cars  may  be  assumed  to  weigh  one  half  that  of  their  contents,  W,  and 
the  frictional  resistances  may  be  allowed  for  at  the  rate  of  50  Ibs.  per  ton 
of  normal  pressure.  Then,  by  formulae  in  Chapter  V, 
13.497=  TVVi2,ooo+fHO(i  +  0.025). 


UNDERGROUND  HAULAGE  SYSTEMS.  293 

W  becomes  79,840  Ibs.,  and  the  hourly  capacity  less  than  160  tons,  without 
allowing  for  delays. 

3.  To  ascertain  if  the  engine  is  properly  proportioned  for  starting  the 
given  load,  we  examine  the  table  on  page   131.     Note  that  the  coefficient 
is  0.3974,  which  multiplied  by  45X(24)2X4  gives  a  minimum  moment  of 
41,202  ft.-lbs.     With  a  direct-acting  engine  the  drum  must  have  a  diameter  of 
5  feet  6  inches.     The  engine  can  therefore  start  from  the  bottom  a  load,  W, 
greater  than  99,300  Ibs. 

4.  If  the  plane  were  double-tracked  and  the  engine  supplied  with  two 
•drums  fast  on  the  same  shaft,  the  only  work  falling  upon  the  engine  would 
be  that  due  to  the  contents  of  the  cars  plus  the  total  friction  on  the  double  line. 
In  this  case  seven  full-length  trips  might  be  made  per  hour,  thus  increasing 
the  hourly  capacity  to  377  tons. 

5.  Required  the  length  of  a  haulage   plant   for  a  tail-rope  system  to 
deliver  icco  tons  in  ten   hours  over  a  road  the  first  600  feet  of  which  has 
a  dip  of  4  feet  per  100  with  the  load,  the  next  2100  feet  a  grade  of  3  feet  per 
100  against  the  load,  and  the  lower  600  feet  2  feet  per  100  against  the  load. 

Assume  that  each  car  weighs  1000  Ibs.,  and  that  its  capacity  is  2000  Ibs.; 
assume  also  a  piston  speed  of  200  feet  per  minute  and  an  average  train-speed 
of  6  miles  an  hour. 

On  the  lowest  section  gravity  produces  a  tension  on  the  main-rope  of 
5665  Ibs.,  and  the  engine  has  to  perform  2,991,120  ft.-lbs.;  on  the  next 
upper  section  for  4  minutes  122.6  horse-power  is  necessary  to  pull  the  train 
up-grade.  At  the  head  of  the  gangway  the  grade  favors  the  loaded  cars, 
which  produce  a  tension  on  the  tail  rope  of  over  6000  Ibs.  With  a  drum 
of  132  inches  diameter,  making  15.3  revo  utions  per  minute,  gearing  to  a 
pinion  of  0.40  its  diameter,  a  modulus  of  |,  and  an  effective  steam  pressure 
of  40  Ibs.,  the  diameter  of  each  cylinder  would  be  21  inches  and  the  stroke 
30  inches. 

REFERENCES. 

Electric  Haulage  in  Metal  Mines,  W.  N.  Clark,  Eng.  &  Min.,  Feb.  25, 
1904;  Benzine  Locomotives  for  Gaseous  Mines,  L.  Volf,  Oesterr.  Zeitschr.  f. . 
Berg.  u.  Huttenwesen,  Oct.  4,  1902;  Electric  Haulage  Plant,  Berwind-White 
Collieries,  T.  W.  Sprague,  E.  &  M.  Jour.,  LIX,  508;  Electric  Mine  Loco- 
motives, William  L.  Affelder,  Third-rail  Friction  Locomotive,  M.  &  M., 
Vol.  XXIV,  383. 

Notes  on  the  Compressed-air  Haulage  Plate  at  No.  6  Colliery  of  the  Sus- 
quehanna  Coal  Co.,  Glen  Lyon,  Pa.,  J.  H.  Bowden,  A.  I.  M.  E.,  Vol.  XXX, 
566. 

Rope  Hauling,  Tail-rope,  Double  Tail-rope,  Pa.  Mine  Insp.,  1889,  340; 
Underground  Wire-rope  Haulage,  Sci.  Am.  Sup.,  Dec.  27,  1902;  Wire-rope 


2Q4  MANUAL   OF  MINIXG. 

Slope  Haulage,  M.  &  M.,  April  1904;  Slope  Haulage,  M.  &  M.,  Vol.  XXIV, 
413;  Rope  Transmission,  M.  &  M.,  Sept.  1900,  89;  Aerial  Ropeways, 
A.  I.  M.  E.,  Vol.  XIX,  760;  Hildbrandt  Essay,  53. 

Engine  Planes  at  Wearmouth  Colliery,  W.  R.  Bell,  Trans.  M.  &  M.,  Eng., 
XLV,  219. 

Report  of  Tail-rope  Committee,  N.  E.  I.,  XVII;  Tail-rope  System,  De- 
scription of  Plant,  R.  S.  Williamson,  Coll.  Guard.,  Nov.  1896,  1096;  Third- 
rail  Tail-rope  Haulage,  M.  &  M.,  XXIII,  465. 

Endless  Rope,  Hildbrandt  Essay,  37;  The  Endless-rope  Haulage  System, 
M.  &  M.,  Vol.  XX,  251,  520,  and  566. 

Brake  for  Inclines,  Coll.  Guard.,  Vol.  LXXXII,  846;  Steam  Locomotives 
for  Mining,  Cassier's  Mag.,  July  1904,  291;  Air  and  Internal  Combustion, 
Aug.;  and  Electric,  Oct.  1904. 

Mules,  Care  of:  Protection  in  the  Mines,  M.  &  M.,  Vol.  XXIII,  272,  568; 
Shoeing  of  Pit  Horses,  J.  A.  Longden,  Brit.  Soc.  Min.  Stud.,  IV,  127;  De- 
scription of  Improved  Colliery  Stahles,  J.  P.  Kirkup,  Brit.  Soc.  Min.  Stud.,  I, 
328;  Selection  and  Care  cf  Horses,  M.  E.  Boissier,  Coll.  Guard.,  June  1897, 

"73- 

Tracks,  Latches,  etc.,  Hildbrandt  Essay,  18,  42;  Self-acting  Switch,  X, 
Coll.  Eng.,  Nov.  1895,  78,  80;  Roads  for  Mines,  James  W.  Abbott,  Eng.  & 
Min.  Jour.,  May  26,  1903;  Grades  in  Rooms,  M.  &  M.,  Vol.  XXII,  227; 
in  Roads,  M.  &  M.,  Vol.  XXII,  180;  Maintenance  of  Track,  M.  &  M., 
Mar.,  199,  338. 

Different  Methods  of  Mine  Haulage  Compared,  B.  F.  Jones,  M.  &  M., 
Aug.  1902;  Comparison  of  Haulage  Methods,  M.  &  M.,  Vol.  XXIII,  8. 

Timbering  in  Mines,  Mineral  Industry,  740. 

The  Underground  Ry.  of  the  Stassfurt  Salt  Mines,  J.  Westphal,  Glauckauf, 
Oct.  4,  1902;  The  Longest  Mine  Haulage,  F.  Z.  Shellenberg,  A.  I.  M.  E., 
Vol.  XXIX,  101. 

Haulage  Tests,  Dynamometer  Truck  Measuring  Resistances,  Coll.  Guard., 
Nov.  20,  1896,  981;  Sheaves,  Pulleys,  Tracks,  etc.,  Hildbrandt  Essay,  38; 
Mechanical  Devices  at  Coal  Mines,  L.  L.  Logan,  M.  &  M.,  Feb.  1904; 
Sheaves,  Pulleys,  Tracks,  etc.,  at  the  Leith  Coal  Mine,  Pa.,  H.  L.  Auchmuty, 
Coll.  Guard.,  Eng.,  Aug.  1896,  6;  Checks  for  Sharp  Curves  on  Engine  Planes, 
Idris  Thomas,  Coll.  Guard.,  April  6,  1900. 

Curves  in  Roads,  M.  &  M.,  Vol.  XXIII,  280,  456;  Track  Switches, 
M.  &  M.,  Vol.  XXIV,  331. 


CHAPTER  IX. 

WIRE-ROPE  TRANSMISSION. 

Ropes  for  Power  Transmission  are  of  different  materials, 
cotton,  hemp,  manilla,  and  steel  wire  being  the  most  common. 
For  mining  purposes  the  cotton  ropes  are  relatively  weak  com- 
pared with  manilla,  and  have  given  place  to  the  latter,  which  are 
still  largely  employed  for  some  purposes  of  driving.  If  care  is 
taken  in  the  proper  selection  of  a  vegetable-fibre  rope  and  in  pro- 
viding it  with  large  pulleys,  its  durability  is  increased  to  an  extent 
that  will  make  it  a  good  competitor  of  wire  rope,  which  is  the 
most  extensively  used  for  power  transmission.  The  steel-wire 
rope  is  of  the  y-wire,  6-strand  type  of  a  diameter  of  f  to  f  inch, 
according  to  the  amount  of  power  desired. 

Rope-driving  Systems. — Motion  may  be  imparted  to  an 
endless  rope  by  its  adhesion  to  a  sheave  around  which  it  passes 
and  thence  to  a  distant  sheave.  If  the  distance  between  the 
sheaves  is  great,  the  rope  is  supported  at  intermediate  points. 
The  amount  of  power  which  can  be  delivered  to  the  far  sheave 
depends  solely  upon  the  amount  of  adhesion  between  the  rope 
and  the  surface  of  the  sheave.  In  the  American  system  one 
endless  rope  is  wound  several  times  around  the  driving  and  the 
driven  pulleys.  In  the  English  system  the  sheaves  have  mul- 
tiple grooves.  Several  independent  ropes  run  side  by  side,  each 
driving  in  its  own  groove.  The  latter,  or  multiple,  system  more 
easily  transfers  power  between  the  generator  and  motor,  and  is 
more  secure  against  the  possibility  of  breakdowns.  On  the 
other  hand,  the  ropes  have  an  unequal  stretch  and  are  liable  to 
drive  at  different  velocities  and  tensions,  thus  introducing  loss 

295 


296  MANUAL  OF  MINING. 

of  power,  greater  wear,  etc.  In  the  American,  or  continuous, 
system  there  is  only  one  splice  in  the  rope  and  no  differential 
driving  is  likely  to  exist. 

The  Transmission  of  Power. — The  amount  of  adhesion  deter- 
mining the  power  transmission  depends  on  the  friction  between 
the  surfaces  in  contact,  the  angle  of  wrap,  and  tensions  on  the 
rope.  Being  of  slow  speed,  the  tensions  in  the  two  branches 
of  the  rope  are  due  solely  to,  and  depend  directly  upon,  the  stretch 
produced  by  the  weight  of  the  sections  between  the  supports  and 
the  amount  of  sag. 

Sheaves  and  End  Carriages. — The  sheaves  around  which 
the  power-rope  travels  are  horizontal,  the  upper  (Fig.  94)  being 
on  a  fixed  frame,  and  the  lower  one  at  the  end,  remote  from  the 
power,  on  a  carriage-tower  frame.  These  sheaves  are  of  large 
diameter  keyed  to  a  heavy  shaft,  the  latter  running  in  a  step  at 
its  foot  and  held  in  a  box  at  the  upper  end.  The  rate  of  revolu- 
tion of  the  sheaves  is  125,  or  less,  per  minute,  the  speed  of  the 
rope  being  maintained  below  4200  feet  per  minute.  The  rim 
of  the  pulley  holds  in  its  groove  a  number  of  wooden  blocks, 
or  a  strip  of  rubber,  to  furnish  as  large  a  friction  as  possible. 
In  some  cases  the  rim  is  composed  of  a  number  of  gripping- 
jaws,  which  are  capable  of  locking  on  the  rope.  The  brake- 
wheel  is  bolted  above  the  arms  of  the  grip-sheave,  and  is  controlled 
by  an  adjusting-screw  and  hand-wheel.  In  power  transmission 
both  frames  carrying  the  end  sheaves  are  fixed,  and  provision  is 
made  for  the  stretch  of  the  rope  by  a  sheave  with  arms  which 
are  capable  of  extension,  thus  enlarging  the  diameter  of  the 
sheave.  The  rope  may  also  be  cut  out  and  respliced. 

At  intermediate  points  on  the  line  sheaves  are  supported 
on  the  pillars  or  towers  to  relieve  the  rope  of  the  excessive  stress 
of  a  long  rope.  These  are  of  cast  iron,  1 8  or  20  inches  in  diame- 
ter, and  fixed  on  the  ends  of  cross-arms.  They  run  loose  on 
their  axles.  The  grooves  are  deep  and  wide  to  admit  the  rope 
nearly  to  the  bottom,  the  sides  being  curved  to  let  the  rope  slip 
easily.  Occasionally  the  grooves  are  lined.  The  distance  be- 
tween the  sheaves  is  governed  by  the  allowable  tension  and  sag. 


WIRE-ROPE   TRANSMISSION.  297 

For  power  purposes  the  maximum  is  450  feet;  for  aerial  tram- 
ways it  may  be  only  200  feet.  If  they  merely  support  the  ropes, 
they  are  small  in  diameter,  but  if  intended  to  deflect  the  rope, 
their  diameter  is  increased.  Care  must  be  taken  in  their  align- 
ment in  a  vertical  plane  and  that  the  axles  each  have  as  little 
friction  as  possible.  On  steep  inclinations  a  second  sheave 
vertically  above  the  lower  supporting  sheave  is  placed  to  prevent 
the  rope  from  flying  from  the  wheel.  Where  it  is  possible  for 
power  transmission,  the  upper  branch  of  the  rope  is  made  the 
driving  portion,  and  the  lower  branch  its  slack  part. 

Maintaining  Uniform  Tension. — All  single-rope  power-trans- 
mission lines  require  some  form  of  tightener  to  maintain  a  con- 
stant tension  upon  the  rope  and  a  uniform  driving  power.  Wire- 
rope  lines  for  endless-rope  haulage  and  aerial  tramways  are 
illustrations  of  this  type.  In  the  English  multiple-rope  system 
tighteners  are  not  used. 

The  tightener  may  be  a  sheave  or  a  pair  of  drums  on  a  mov- 
able carriage.  When  the  drums  are  at  the  power  end  of  the 
line  they  are  placed  on  the  same  base  with  their  axes  parallel. 
The  rope  is  wound  consecutively  from  one  to  the  other  of  the 
drums  and  around  all  of  the  grooves.  Either  or  both  may  then 
act  as  the  driver  if  they  are  geared  together.  The  two  may  also 
be  mounted  on  a  carriage,  in  which  case  they  are  at  the  driven 
end  of  the  line.  The  carriage  is  free  to  travel  upon  a  track  and 
is  fitted  with  a  heavy  pendent  weight,  such  that  when  the  rope 
stretches  the  counterpoise  will  pull  the  carriage  to  balance  the 
tension,  as  in  Fig.  93.  The  length  of  the  travel  of  the  carriage 
must  be  sufficient  to  allow  of  a  freedom  of  movement  within 
limits  of  variation  in  rope  length  caused  by  the  changes  of  load. 

The  Tension  of  the  Rope. — On  the  tight  side  the  tension  is 
that  which  is  due  to  the  centrifugal  action  of  the  pulley-sheave, 
the  bending  stress  of  the  rope,  and  the  power  which  the  latter 
receives  from,  or  remits  to,  the  respective  sheave.  This  must 
never  exceed  the  maximum  safe  stress  of  the  rope  and  should 
be  less  than  one  third  of  the  ultimate  strength.  The  tension  on 
the  slack  side  is  equal  to  that  on  the  driving  side  less  the  power 


-298  MANUAL  OF  MINING. 

taken  off  by  the  sheave.  It  is  from  a  half  to  one  quarter  of  the 
tension  on  the  driving  side,  and  the  relation  between  these  three 
forces  is  indicated  by  the  following  formulae: 

Let  v=  velocity  in  feet  per  second; 
d=  diameter  of  the  rope,  inches; 
R  =  weight  of  the  rope  per  foot; 
Z=distance  between  stations,  feet; 
TT=weight  on  the  journals; 
^^revolutions  per  minute; 
S=  maximum  safe  stress  of  the  rope; 
k=  bending  stress  of  the  rope; 

7\=  available  working  tension  of  the  rope,  maximum; 
C  =  stress  due  to  centrifugal  force; 
T=  tension  on  the  slack  side  of  the  rope; 
P=  power  transmitted  to  the  pulley,  Ibs.;  and 
H=  horse-power  transmitted. 

Then 
C=o.i3^2;     Tt=S-C-k;    P  =  T1-T;     and     H  =0.00182  Pv. 

The  influence  of  centrifugal  force  is  negligible  in  power  trans- 
missions for  underground  haulage  or  aerial  tramways.  In  these, 
too,  the  difference  in  tensions  is  not  great,  for  the  dead  load  of 
cars,  buckets,  or  conveyors  is  exerted  nearly  equally  upon  both 
branches  of  the  line. 

The  Sag  of  a  Rope. — The  tension  upon  the  rope  may  be 
regulated  by  the  amount  of  sag,  or  deflection,  permitted  between 
the  two  points  of  support.  The  relation  between  them  is  indi- 
cated by  the  formula  below.  Having  calculated  the  sag  for  a 
given  power  condition,  a  mean  between  h  and  At  is  taken.  The 
end  sheaves  are  separated  sufficiently  to  reduce  the  sag  at  rest 
to  this  amount.  When  in  action  the  desired  tensions  will  be 
obtained.  If  more  power  is  desired,  the  sheaves  are  separated 
somewhat.  If  the  desired  amount  cannot  be  obtained,  the  rope 
is  cut  and  spliced.  This  must  be  done  every  three  months,  as 
wear  and  use  weaken  the  rope. 


WIRE-ROPE   TRANSMISSION.  299 

The  deflection  of  a  power-  rope  at  the  centre  of  its  span  in 
feet  being  h  for  the  slack  and  h^  for  the  tight  portion,  then 

RV  RL2 

h  and    h= 


Where  possible,  the  upper  branch  of  the  rope  should  be  the 
driver  and  the  lower  branch  the  slack  portion.  An  average 
sag  on  spans  of  150  to  250  feet  is  about  3  feet  when  the  rope  is 
at  rest,  this  amount  being  about  one  half  of  the  sag  permitted 
for  the  slack  portion  when  in  action. 

EXAMPLE.  —  Power  is  to  be  transmitted  by  a  steel  7-wire  rope,  i  inch  in 
diameter,  at  a  velocity  of  4200  feet  per  minute  over  i2-ft.  wooden  grooved 
pulleys  400  feet  apart.  Required  the  horse-power  which  can  be  trans- 
mitted and  the  deflections  to  be  given  the  rope. 

Let  5=2i,ooo  Ibs.  for  the  rope,  and  assume  the  tension  on  the  slack 
side  to  be  one  half  that  on  the  tight  side. 

Here  ^=70;    ^=0.5^;  and  £=1.58. 

Then  C=o.i3^2=637;  £=10,073  Ibs.;  7^=10,290;  ^=0.5^  =  5145; 
P=5i45  Ibs.;  #=646.47  horse-power  as  a  maximum;  ^  =  8.2  feet;  and 
£=4.1  feet. 

Aerial  Tramways.  —  Elevated  endless  ropes  are  a  special  form 
of  power  transmission.  They  have  long  been  used  as  a  cheap 
mode  of  transportation,  serving  as  feeders  to  establish  systems 
of  communication.  They  consist  of  a  ropeway  elevated  above 
the  surface,  with  carriers  for  mineral  swinging  free  above  the 
ground.  The  elevated  structure  is  essentially  a  series  of  pillars, 
or  towers,  of  a  strength  depending  upon  the  load  and  its  speed 
and  the  inclination.  It  is  employed  where  the  contour  is  too 
rugged  for  economical  road-building,  and  is  therefore  attractive 
to  the  miner  and  the  quarryman.  The  mineral  is  charged  into 
the  carriers  at  the  mine  and  these  are  allowed  to  descend,  tak- 
ing with  them  the  rope  to  which  they  are  attached  if  the  grade 
is  sufficient,  or  are  pulled  downward  if  power  must  be  employed. 
A  succession  of  carriers  is  loaded  at  the  top  and  delivers  the 
contents  at  the  bottom,  where  they  are  automatically  emptied 
without  stop,  and  thence  returned  on  the  empty  line.  The  rope 


300  MANUAL  OF  MINING. 

may  be  straight  and  of  uniform  grade,  or  follow  the  surface  con- 
figuration over  undulating  profile  and  around  curves  for  dis- 
tances of  a  mile  or  more,  provided  the  power  be  sufficient  to 
overcome  the  resistance. 

The  power  required  to  operate  aerial  roadways  depends  upon 
the  grade  and  average  inclination  of  the  line.  In  the  simplest 
case  the  grade  may  be  uniform  to  the  point  of  shipment.  Then 
the  inclination  alone  will  determine  if  auxiliary  power  is  necessary 
either  to  assist  the  load  or  to  check  its  speed.  If  the  inclination 
exceeds  20  per  cent,  the  system  is  self-acting  and  only  a  slight 
braking  power  is  requisite  at  the  top  of  the  line.  If  it  exceeds 
this  amount,  power  is  required  to  check  the  speed.  The  brake 
must  overcome  the  excess  of  the  downward  effort  of  the  load 
over  the  resistance  of  the  rising  empty  bucket  and  the  total 
friction.  If  the  grade  is  below  20  per  cent,  power  may  be  re- 
quired to  supplement  the  force  of  gravity.  It  is  not  necessary 
that  the  grade  be  uniform,  provided  there  is  ample  fall  from  the 
mine  to  the  roadway.  In  this  event  great  care  must  be  exer- 
cised in  determining  the  value  of  the  motor  force  as  well  as  all 
the  elements  which  contribute  to  the  resistance.  If  power  is 
required  in  such  a  case,  it  may  be  applied  at  either  end  of  the 
line  upon  a  grip-sheave.  The  motor  attached  or  belted  to 
the  driving- sheave  must  be  of  the  slow- speed  type.  The  super- 
structure is  light  and  inexpensive.  There  is  no  interference  with 
or  from  surface  travel.  Climatic  conditions  do  not  interrupt 
operations.  The  time  for  their  construction  is  comparatively 
short,  and  they  have  a  capacity  equal  to  that  of  the  average  mine 
usually  located  in  such  districts.  There  is  no  machinery  along 
the  line  except  the  towers  with  their  sheaves.  In  elevated  rope- 
ways only  such  driving  and  controlling  apparatus  is  built  as 
may  be  required  for  maintaining  a  constant  speed  and  uniform 
tension.  The  expense  of  operation  is  comparatively  low. 

The  Single-rope  Tramway. — The  earliest  of  these  tramways 
was  a  single  endless  moving  rope,  of  which  the  Hodgson  and 
Hallidie  are  the  representatives.  The  buckets  or  the  carriages 
are  attached  to  saddles  which  ride  on  the  rope.  They  are  either 


WIRE-ROPE  TRANSMISSION. 


301- 


attached  permanently  to  the  rope  or  may  be  adjustable.     In  both 
these  patterns  one  and  the  same  rope  moves  the  load  (Fig.  117),- 


The  Ore-carriers. — The  mineral  is  transported  in  buckets? 
of  various  designs,  according  to  the  character  of  the  material  to 
be  handled.  They  are  suspended  by  hangers  or  clips,  which  are 
either  inserted  into  the  rope  or  clinched  around  the  outside  of  it,, 
and  attached  at  intervals  determined  by  the  amount  of  material 


3C2 


MANUAL  OF  MINING 


FIG.  118. — The  Power  End  of  the  Halliclie  Svstem. 


FIG.  119. — The  Ore-buckets  of  a  Single-rope  Tramway. 


WIRE-ROPE   TRANSMISSION.  303? 

to  be  delivered.  Usually  they  are  wrought-iron  rectangular 
buckets  holding  about  TOO  Ibs.  each  (Fig.  119).  For  the  trans- 
port of  very  large  outputs  the  buckets  may  be  nearer  together 
than  the  average  200  feet,  or  larger,  and  the  rope  may  be  heavier 
than  the  ordinary  size  of  f  inch.  The  buckets  may  be  loaded  at 
any  point  along  the  line,  automatically  or  by  hand.  They  are  un- 
loaded at  the  lower  end  by  an  automatic  device  in  which  the 
carrier  strikes  a  lever,  which  opens  a  catch,  releasing  the  swinging; 
bottom  and  discharging  the  ore.  The  bottom  is  returned  to 
place  by  a  heavy  counterpoise  weight  on  the  end  of  a  projecting; 
arm  and  is  automatically  locked.  The  hangers  are  so  made 
that  they  may  pass  uninterruptedly  over  the  rims  of  the  support- 
ing sheaves  and  around  the  terminal  pulleys. 

Capacity. — A  ropeway  running  200  feet  per  minute,  cany- 
ing  100  Ibs.  per  bucket  every  100  feet,  will  deliver  60  tons  per 
shift.  With  a  descent  sufficient  for  gravity  to  supply  the  power,, 
three  men  can  manage  all  of  its  operations.  It  requires  some' 
supervision,  and  delivers  ore  at  20  to  35  cents  per  ton-mile  (in- 
clusive of  all  allowances),  and  about  60  cents  per  cord-mile  for 
wood.  The  line  can  be  completed  for  $1.30  per  foot,  and  $2000 
for  the  machinery  at  the  terminals.  Curves  and  long  stretches 
increase  the  first  cost;  grade  does  not.  The  greatest  item  in 
the  maintenance  cost  is  that  for  renewals  of  the  clip-hangers. 
It  amounts  to  $100  a  year  on  a  line  of  half  a  mile  in  length. 

The  Cable  Supports. — These  are  simple  frames  or  towers- 
built  up  from  a  rectangular  base.  The  former  construction  is 
rigid  and  not  liable  to  get  out  of  line,  nor  can  its  cross-arm  be 
pulled  down  on  the  loaded  side  of  the  rope.  In  Figs.  117  and 
1 20  are  two  typical  constructions.  The  height  of  the  stations 
depends  upon  the  distance  between  them.  It  is  about  20  feet 
when  the  stations  are  200  feet  apart.  The  frames  are  locatei 
at  the  higher  points  always  when  there  are  any  undulations  irr 
the  grade  of  the  hill,  as  shorter  towers  can  be  used  with  equal 
result  and  the  inclination  of  the  rope  is  more  nearly  natural. 
On  top  of  each  is  bolted  a  cross-arm,  at  the  ends  of  which  are 
boxed  the  carry  ing- sheaves  to  support  the  rope  and  allow  free- 


304  MANUAL   OF  MINING. 

•dom  of  movement.  In  some  instances  two  sheaves  are  carried 
an  upper  and  a  lower  one.  The  object  of  the  upper  one  is  tc 
;prevent  the  rope  jumping  out  of  place  from  its  groove  in  the 
jlower.  At  a  curve  the  standards  are  near  together  and  the  rope 
jis  slightly  deflected  at  each  standard  to  conform  to  the  curve 
-desired.  The  sharp  turns  or  horizontal  angles,  as,  for  example, 
.around  bluffs,  are  also  made  by  using  large  sheaves.  One  only 
jis  necessary  for  the  turn  when  the  clips  holding  the  buckets  pro- 
ject outward  from  the  rope,  but  two  will  be  required  when  the 
clips  project  inward.  In  the  latter  case  the  ropes  cross  in  pass- 
ing to  the  first  pulley  and  also  to  and  from  the  second  pulley. 

The  end  sheaves  and  carriages  are  similar  to  those  described 
for  power  transmission,  with  modifications  dependent  upon  local 
conditions.  In  Fig.  118  is  illustrated  the  upper  frame  of  the  line, 
while  the  lower  end  is  illustrated  in  Fig.  117,  on  the  left. 

The  Bleichert  System. — In  the  two-rope  system  there  is  a 
pair  of  separate  stationary  cables  over  which  numerous  trolley 
carriages,  supporting  buckets,  are  drawn  by  a  light  endless  mov- 
ing  rope,  called  the  traction-rope,  to  which  the  buckets  are 
Attached  by  patent  grip  (Fig.  120).  The  cables  constitute  the 
roadway  for  the  trolleys.  The  tubs  are  dumped  automatically 
into  a  bin  or  wagon  at  the  lower  end  of  the  line. 

The  several  varieties  of  two-rope  systems  differ  mainly  in 
the  mode  of  suspending  the  rope.  In  one  system  the  bucket  is 
.attached  to  a  clutch,  which  seizes  the  running  rope  and  rails  and 
attaches  by  friction  alone.  It  locks  and  releases  automatically, 
.and  exerts  a  uniform  friction  whatever  may  be  the  slope  of  the 
ropes.  The  rope  permits  them  to  be  disconnected  and  run  off 
•  on  suspended  rails  when  they  reach  their  terminal,  or  any  turn- 
out, on  the  line.  This  enables  them  to  be  loaded  and  discharged 
without  interference  with  the  travel. 

The  double- rope  system  has  a  greater  capacity  than  the 
;  single  rope.  It  is  capable  of  much  heavier  traffic,  for  the  indi- 
vidual loads  may  be  as  great  as  1000  Ibs.  each,  and  the  distance 
between  the  towers  1000  feet,  though  the  average  span  is  300 
.feet.  The  individual  loads  carried  by  the  Hallidie  lines  cannot 


WIRE-ROPE  TRANSMISSION. 


3°S 


exceed  150  Ibs.,  nor  the  distance  between  the  towers  300  feet. 
The  buckets  cannot  be  closer  than  the  distance  between  the  towers. 
The  speed  of  the  rope  is  limited  to  300  feet  per  minute,  because 
of  the  danger  of  the  rope  jumping  off  the  carriage- sheaves. 

TRAMWAY  CAR  FOR  THE  TRANSPORTAffON  -OF  GOAL.  ORES.  SANDS,  fcc. 
SHOWING   LUG  COUPLING 


FIG.  120. — A  Double-rope  Aerial  Roadway. 

The  operating  cost  of  the  double- rope  system  is  higher  than 
that  of  the  Hallidie  single  ropeway,  but  its  capacity  is  also  larger. 
The  materials,  and  supplies,  in  one  instance,  for  a  two-rope  sys- 
tem was  $1.90  per  foot,  exclusive  of  the  machinery.  The  net 
cost  of  a  single  ropeway  in  the  immediate  vicinity  of  the  above 
is  $1.21  per  foot  of  line,  plus  the  cost  of  the  two  end  structures. 
The  cost  of  repair  of  either  system  is  not  heavy,  but  the  double- 
rope  system  is  less  than  that  of  the  single- rope.  On  one  of  the 
former,  two  miles  in  length,  the  cost  of  maintenance  is  1.5  cents 
per  ton  put  of  the  total  operating  cost  of  17.5  cents  per  ton. 

A  self-acting  aerial  tramway,  like  the  self-acting  underground 
plane,  may,  by  stretching  one  or  two  ropes  the  full  length  of 
the  line,  serve  as  a  guide  for  large  skips  holding  a  ton  or  so 
which  are  attached  to  an  endless  rope.  The  latter  rope  is  of  a 
length  equal  to  that  of  the  roadway,  and  is  operated  from  a  clip- 
pulley  at  the  top  by  gravity  with  or  without  power,  according  as 
the  grade  is  small  or  large. 


306  MANUAL  OF  MINING. 


REFERENCES. 

Aerial  Tramways,  at  Sandor,  C.  T.  Finlayson,  B.  C.  Min.  Sci.  Press,  June 
1897,  544;  Various  Types  of  Aerial  Ropeways,  W.  Carrington,  Coll.  Guard., 
Mar.  1897, 556;  Wire-rope  Tramways,  E.  &  M.  Jour.,  LXI,  208;  The  Trans- 
mission of  Power  by  Wire  Rope,  W.  H.  Graves,  M.  &  M.,  April  1904;  Wire- 
rope  Tramways,  J.  H-  Janeway,  Jr.,  M.  &  M.,  April  1904;  Aerial  Ropeways, 
A.  I.  M.  E.,  Vol.  XIX,  760;  Hildbrandt  Essay,  53. 

New  Rope  Clip,  Coll.  Guard.,  Vol.  LXXX,  823. 

Compression  of  Air,  Peabody,  Thermodynamics,  Coll.  Eng.,  XVII,  173; 
Air  Compressors,  Valve  Movements,  Coll.  Eng.,  Feb.  1897,  319;  A  Study  in 
the  Economical  Arrangement  of  Compressed-air  Haulage,  E.  &  M.  J.,  Feb. 
28, 1903.  i 


CHAPTER  X. 

THE  COMPRESSION  OF  AIR. 

Compressed  Air  is  employed  in  the  same  manner  as  steam  for 
motor  purposes  in  driving  rock-drills,  pumps,  locomotives,  and 
coal-cutters.  The  advantage  over  steam  lies  in  its  transmission 
without  condensation  and  giving  cool,  dry,  ventilated  rooms 
instead  of  hot  rooms  from  exhaust  where  steam  is  used.  It  is, 
however,  dearer  than  the  other  motor  agency.  It  can  be  trans- 
mitted over  great  distances  where  electric  transmission  is  not 
desirable  or  economical. 

When  air  is  subjected  to  pressure  its  volume  is  proportion- 
ately diminished,  and  the  energy  thus  expended  in  the  compression 
is  retained  by  the  air  and  capable  of  being  applied  as  is  the  ten- 


FIG.  121. — A  Water-jacketed  Air-compressor. 

sion  of  steam.  To  obtain  air  at  an  absolute  pressure  of  100  Ibs. 
per  square  inch  a  unit  volume  must  be  reduced  to  0.147;  a^  a 
pressure  of  200  Ibs.  per  square  inch,  to  0.074.  In  order  to  obtain 
the  same  pressure  from  steam  it  must  be  heated  to  338°  F.  and 
388°  F.  respectively. 

The  air-compressor  is  an  ordinary  cylinder  provided  with  a 
piston  and  suitable  valves  for  admitting  and  delivering  air.    During 

307 


308 


MANUAL  OF  MINING. 


the  out-stroke  air  flows  into  the  cylinder  through  the  inlet- valves 
AA,  Fig.  121,  which  close  when  the  return-stroke  commences. 
The  air  in  the  cylinder  is  then  compressed  until  it  opens  the 
delivery-valves,  BB,  through  which  it  is  forced  into  a  receiver. 

The  Weight  of  Air. — The  following  table  of  weights  of  a 
cubic  foot  of  air  at  different  temperatures  has  been  calculated 
on  the  assumption  of  a  constant  pressure  of  30  inches  of  mercury, 
approximately  sea-level  pressure: 


Tempera- 
ture /. 

Weight  in 
Decimals  of 
a  Pound. 

Tempera- 
ture *. 

Weight  in 
Decimals  of 
a  Pound. 

32 

.0809749 

120 

.0686678 

35 

.0804831 

125 

.0680799 

40 

.0796767 

130 

.0675020 

45 

.0788863 

135 

.0669338 

5° 

.0781113 

140 

.0663751 

55 

•0773515 

145 

.0658256 

60 

.0766063 

IS0 

.0652852 

62 

.0763122 

155 

•0647535 

65 

•0758753 

160 

.064^305 

70 

.0751582 

165 

.0637158 

75 

.0744544 

170 

.0632093 

80 

•0737638 

175 

.0627108 

85 

.0730858 

180 

.0622201 

QO 

.0724202 

185 

.0617371 

95 

.0717666 

190 

.0612614 

100 

.0711246 

195 

.0607931 

105 

.0704941 

200 

.0603318 

no 

.0698746 

205 

•0598775 

"5 

.  0692660 

212 

.0592529 

The  table  on  page  309  shows  the  weight  of  air  at  various 
altitudes  above  the  sea. 

Free  Air. — Free  air  is  understood  as  air  at  the  ordinary  tem- 
perature and  an  atmospheric  pressure  of  14.7  Ibs.  per  square 
inch  or  2116.8  Ibs.  absolute  per  square  foot.  By  "absolute  pres- 
sure" is  meant  the  pressure,  above  the  vacuum  as  distinguished 
from  gauge  pressure,  which  is  measured  above  the  atmosphere 
by  the  gauge.  Absolute  temperature,  at  Fahrenheit  scale,  is 
the  reading  of  the  thermometer,  in  Fahrenheit,  plus  461°, 
and  is  represented  by  T.  At— 461°  F.  there  is  no  pressure. 
The  absolute  temperature  is  o°. 


THE  COMPRESSION   OF  AIR. 


3°9 


TABLE  OF  ABSOLUTE  PRESSURES,  BOILING-POINTS,  ETC.,  AT  DIFFERENT 
HEIGHTS  ABOVE  SEA-LEVEL. 


I 

• 

3 

4 

s 

6 

7 

Height 
above 
Sea-level, 
Feet. 

Barometer, 
Inches  of 
Mercury. 

Boiling- 
point, 
Degrees 
Fahr. 

Absolute 
Pressure, 
Pounds. 

Weight  of 
i  Cubic 
Foot  of 
Air  at  60°, 
Pounds. 

Volume  of 
Air  Equal 
to  i  Cubic 
Foot  of 
Free  Air  at 

Volume  of 
Free  Air  at 
Sea-level 
Equal  to  i 
Cubic  Foot 
at  Given 

Sea-level. 

Altitude. 

0 

30 

212 

14-7 

.0765 

I 

I 

5" 

29.42 

211 

14.41 

•  07499 

.02 

.98039 

1,025 

28.85 

210 

14.136 

•07356 

.04 

.96154 

^539 

28.29 

209 

13.86 

.07213 

.06 

•9434 

2,063 

27-73 

208 

I3-587 

.07071 

.08 

•9259 

2,589 

27.18 

207 

I3-3I8 

.0693 

.10 

.90909 

3>IT5 

26.64 

206 

I3-054 

.06793 

.12 

.89285 

3,642 

26.  II 

205 

12.794 

.06658 

.14 

.87719 

4,169 

25-59 

204 

12-539 

-06525 

•17 

•8547 

4,697 

25.08 

203 

12.289 

•06395 

.I9 

.8403 

5,225 

24-58 

2O2 

12.044 

.06267 

.  22 

.8197 

5,764 

24.08 

2OI 

i  i  .  799 

.0614 

.24 

.8064 

6,304 

23-59 

2OO 

11  -559 

.06015 

•27 

•7874 

6,843 

23.11 

199 

11.324 

•05893 

•29 

•7752 

7,38i 

22.64 

I98 

11.094 

•05773 

•32 

•75757 

7,932 

22.17 

197 

10.863 

•0565 

•35 

•  74074 

8,481 

21.71 

196 

10.638 

•05536 

•38 

.7246 

9,031 

21  .  26 

195 

10.417 

.05421 

•4i 

.7092 

9,579 

20.82 

194 

10.202 

•05309 

•44 

•  6944 

0,127 

20-39 

193 

9-99 

.05199 

•47 

.6802 

0,685 

19.96 

192 

9.78 

.0509 

•50 

.6666 

1,243 

J9-54 

I9I 

9-57 

.0498 

•53 

•6536 

T,799 

!9-!3 

I90 

9-37 

.0488 

-56 

.64102 

2,367 

18.72 

ISO 

9.17 

.0477 

.60 

.625 

2,934 

18.32 

188 

8.98 

.0467 

•63 

•6i35 

13,498 

17-93 

187 

8.78 

•0457 

•67 

.6 

14,075 

17-54 

186 

8-59 

.0447 

•71 

.5848 

14,649 

17.16 

185 

8.41 

•°437 

•74 

•5747 

The  volume  of  air  may  be  altered  by  a  change  of  temperature 
•or  of  pressure.  It  increases  with  a  rise  in  temperature  or  a 
reduction  of  pressure.  Correspondingly  it  decreases  for  a  di- 
minished temperature  or  an  increased  pressure. 

Adiabatic  Compression. — When  compressed  or  expanded  by 
•the  application  of  force,  air  suffers  a  change  of  temperature.  If 
the  walls  confining  the  air  be  absolute  non-conductors  of  heat, 
the  increase  in  temperature  and  of  pressure  may  be  definitely 
known. 


310  MANUAL  OF  MINING. 

The  relation  between  the  pressure,  volume,  and  temperature 
is  expressed  by  the  following  equations,  in  which  P  =  the  initial 
pressure  of  the  air  in  pounds  per  square  foot;  F  =  the  volume  at 
that  pressure  in  cubic  feet;  and  r  =  its  absolute  temperature  P,  V, 
and  T  correspondingly  represent  the  pressure,  volume,  and  tem- 
perature at  the  end  of  the  operation. 


and 


log  Tt  =  log  7+0.29  log  Pj-o.29  log  P; 
log  T,  =log  T+  0.408  log  V—  0.408  log  Fjj 
log  P  +  1.408  log  7  =  log  Pt+  1.408  log  Fx. 

Thus,  if  a  pound  of  free  air  at  60°  be  compressed  to  90  Ibs.  abso- 
lute, it  will  attain  a  final  temperature  of  881°  absolute  or  420°  F., 
the  final  volume  becoming  0.276  of  the  original.  This  is  known 
as  adiabatic  compression. 

log  ^  =  2.716838+0.29X1.954243-0-29X1.176091, 

whence  Tt  =  88i. 

Isothermal  Compression.  —  If  the  air  be  compressed  within 
walls  which  are  perfect  conductors  and  the  compression  be  a 
slow  one,  the  heat  generated  will  be  absorbed  or  radiated  as 
fast  as  developed  and  the  air,  originally  at  60°,  will  remain  at  60°, 
the  process  being  isothermal.  At  the  final  pressure  of  90  Ibs. 
absolute  the  volume  would  be  only  0.1640  of  the  original.  Thus 
isothermal  compression  furnishes  a  smaller  volume  at  a  lower 
temperature  than  is  attained  in  adiabatic  compression. 

The  relation  of  volume,  temperature,  and  pressure  is  repre- 
sented by  the  equation  PF  =  53.i8r.  Thus  the  volume  of  air 
varies  inversely  as  the  pressure  when  the  temperature  is  con- 
stant; the  absolute  pressure  varies  directly  with  an  absolute 
temperature  if  the  volume  remains  constant;  and  the  volume 
varies  as  the  absolute  temperature  so  long  as  the  pressure  re- 
mains constant. 


THE  COMPRESSION   OF  AIR. 


TABLE  OF  VOLUMES,  MEAN  PRESSURES,  TEMPERATURES,  ETC.,  IN  THE  OPERA- 
TION OF  AIR-COMPRESSION  FROM  i  ATMOSPHERE  AND  60°  FAHR.* 


•I 

2 

3 

4 

5 

6 

7 

8 

9 

10 

ii 

, 

*,  ^ 

a 

tc-M 

+> 

M'~g 

«_£ 

• 

o 

I 

^ 

0.5 

fc 

£ 

i 

g 

<3  s 

j« 

o  a 

QO 

Q      T3 

rt    • 

i 

I 

4 

H 

"< 

l*i 

^ 

v      C 

&1 

i 

I 

I 

c 

f  4» 

+3 

„ 

l.ll 

1JJ 

IcS 

1 

ii 

.1 

gj 

5jfl 

•H-d 

0£ 

!!<§| 

III 

-1 

1 

1 

W    <£ 

E-S 

|!l 

3"8 

ocj 

«MH 

-8 

a  cflO 

ij-jjjl 

§  ^  u 

•~'3 

1 

o 

^ 

£ 

^ 

« 

^ 

s 

fe 

o 

0 

14.7 

i 

! 

i 

0 

0 

0 

0 

60 

i 

15-7 

i.  068 

•9363 

•95 

.96 

•975 

•43 

•44 

71 

I 

2 

16.7 

1.136 

.8803 

.91 

1.87 

1  .01 

.96 

.96 

2 

3 

17.7 

1.204 

•8305 

.876 

2.72 

2.8 

1.4 

1.41 

88^9 

3 

4 

i8.7 

1.272 

.7861 

.84 

3-53 

3.67 

1.84 

1.86 

98 

4 

5 

19.7 

i-34 

.7462 

.81 

4-3 

4-5 

2.22 

2.26 

1  06 

5 

10 

24.7 

1.68 

•5952 

.69 

7.62 

8.27 

4.14 

4.26 

145 

10 

15 

29.7 

2.02 

•495 

.606 

10.33 

11.51 

5-77 

5-99 

J5 

20 

20 

34.7 

2.36 

•4237 

•543 

12  .  62 

14.4 

7.2 

7.58 

207 

25 

39-7 

2-7 

•3703 

•494 

J4-59 

17.01 

8.49 

9-°5 

234 

25 

30 

44.7 

3-04 

.3289 

•4638 

16.34 

19.4 

9.66 

10.39 

255 

3° 

35 

49-7 

3-381 

•2957 

.42 

17.92 

21.6 

10.72 

n-59 

281 

35 

40 

54-7 

3.721 

.2687 

•393 

19.32 

23.66 

11.7 

12.8 

302 

40 

45 

59-7 

4.061 

.  2462 

•37 

20.52 

25-59 

12.62 

13-95 

321 

45 

50 

64.7 

4.401 

.  2272 

•35 

21.79 

27-39 

13.48 

15-05 

339 

5° 

55 

69.7 

4-749 

.  2109 

•331 

22.77 

29.  II 

14-3 

15.98 

357 

55 

60 

74-7 

5.081 

.1968 

23.84 

30-75 

16.89 

375 

60 

65 

79-7 

5.423 

.1844 

.301 

24.77 

31.69 

J5-76 

17.88 

389 

65 

70 

84-7 

5.762 

•I735 

.288 

26 

33-73 

16.43 

18.74 

405 

70 

75 

89.7 

6.102 

.1639 

.276 

26.65 

35-23 

17.09 

19-54 

420 

75 

80 

85 

94-7 
99-7 

6.442 
6.782 

•1552 
.1474 

.267 
.2566 

27-33 
28.05 

36.6 

37-94 

17.7 
18.3 

20.5 

21  .  22 

432 
447 

80 

85 

90 

104.7 

7-122 

.1404 

.248 

28.78 

39.18 

18.87 

22 

459 

90 

95 

109.7 

7.462 

•134 

.24 

29-53 

40.4 

19.4 

22.77 

472 

95 

100 

114.7 

7.802 

.1281 

•232 

30.07 

41.6 

19.92 

23-43 

485 

too 

105 

119.7 

•8.142 

.1228 

.2254 

30.81 

42.78 

20.43 

24.17 

496 

105 

no 

124.7 

8.483 

.1178 

.2189 

3T.39 

43  -91 

20.9 

24.85 

5°7 

i10 

"S 

129.7 

8.823 

•"33 

.2129 

31.98 

44.98 

21.39 

25-54 

518 

ii5 

120 

134-7 

9.163 

.  1091 

•2073 

32  -  54 

46.04 

21.84 

26.2 

529 

rao 

125 

139.7 

9-503 

.1052 

.202 

33-07 

47-06 

22.26 

26.8l 

540 

125 

130 

144.7 

9-843 

.1015 

.1969 

33-57 

48  ..i 

22.69 

27.42 

550 

i3° 

135 

149.7 

10.183 

.0981 

.1922 

34  •  °5 

49-i 

23.08 

28.  oe 

1  35 

140 

154-7 

IO-523 

•095 

.1878 

34-57 

50.02 

23-41 

28.66 

57o 

i4° 

145 

159-7 

10.864 

.0921 

•1837 

35  •  °9 

51 

23-97 

29.26 

580 

i45 

164.7 

i  i  .  204 

.0892 

.1796 

35-48 

51.89 

24.28 

29.82 

589 

1  5° 

1  00 

J74-7 

11.88 

.0841 

.1722 

36.29 

24-97 

30.91 

607 

Z6o 

170 

184.7 

12.56 

.0796 

•l657 

37-2 

55-39 

25-71 

32.07 

624 

1  7° 

1  80 

194.7 

13-24 

•°755 

•Z595 

37-96 

57-01 

26.36 

33-04 

640 

j8o 

I90 

204.7 

13.92 

.0718 

•154 

38.68 

58.57 

27.02 

34.06 

6^7 

1  90 

2OO 

214.7 

14-6 

.0685 

.149 

39-42 

60.  14 

27.71 

35-02 

672 

2oo 

*From  Richards'  "Compressed  Air,"  page  21. 


312  MANUAL  OF  MINING. 

The  table  on  page  311  shows  the  relation  between  volume, 
pressure,  and  temperature  in  adiabatic  and  isothermal  com- 
pression. 

The  columns  4,  8,  and  6  respectively  express  the  relations 
assumed  by  one  unit  of  air  when  compressed  without  change 
of  temperature  to  the  stated  pressures.  Column  8  shows  the 
mean  effective  resistance  during  the  period  of  compression  which 
is  offered  by  the  air  to  the  point  of  delivery  only.  Column  5 
contains  the  ratio  of  volumes  of  air  compressed  adiabatically  to 
given  pressures;  column  9,  the  mean  effective  resistance  during 
the  compression  period;  and  column  6,  the  mean  effective 
resistance  during  the  entire  stroke,  including  the  period  of 
delivery. 

EXAMPLES. — i.  Let  it  be  desired  to  determine  the  final  volume  of  i  Ib.  of 
free  air  at  a  temperature  of  62°  F.  compressed  to  80  Ibs.  gauge. 
Then 

V=  13.08;   P  is  14.7X144=2116.8;   P,=94-7;   and    1=62+461  =  523°. 

Tsothermally, 

2ii6.8X  13.08=  27,729=P1F1.    ^,=94.7X144X7,. 
Whence 

V  =  2.027  cu-  ft- 
Adiabatically, 

2ii6.8x(i3.oS)1-408=63,iio=P,(Fl1-408;   94.7Xi44X(Fl)I'408=63,no. 
Whence 

F,- 3.492. 

According  to  the  preceding  table  the  volumes  are  respectively  (columns 
4  and  5)  0.1552X13.08=2.027  cu.  ft.  and  0.267X13.08=3.49  cu.  ft. 
What  is  the  final  temperature  in  the  latter  case? 

log  r=log  523+0.29  log  94.7-0.29  log  14.7=2.953151. 
r=897°abs.  =436°F. 

According  to  the  table,  column  10,  r  would  have  been  432°  F.  from  60°  F. 
2.  Required  the    temperature  of   a  pound  of  air  which   at  50  Ibs.  abs. 
pressure  occupies  a  volume  of  10  cu.  ft. 

PF=5oXi44Xio=53.i8r; 
whence 

T=  1353°  absolute=  892°  F. 


THE  COMPRESSION   OF  AIR.  313 

Air  Indicator  Cards.  —  In  Fig.  122  are  indicator  cards 
steam-  and  air-cylinders;  they  show  also  the  adiabatic  and  iso- 
thermal compression.  On  the  horizontal  the  volumes  are  meas- 
ured from  the  right,  and  on  the  vertical  lines  the  corresponding 
pressures.  It  is  noticeable  that  the  adiabatic  curve  rises  more 
rapidly  than  the  isothermal;  in  other  words,  its  pressure  in- 
creases more  rapidly.  Again,  the  volumes  in  the  former  case  are 
correspondingly  larger  for  a  given  pressure,  as  may  be  seen  by 
comparing  volumes  obtained  by  cutting  the  two  curves  by  some 
horizontal  line.  This  may  also  be  noted  by  comparing  the  vol- 
umes in  columns  5  and  4  of  table  page  311. 


FlG.  122. — Indicator  Cards  from  Steam-  and  Air-cylinders. 

The  initial  pressure  of  the  steam  in  Fig.  122  is  58  Ibs.  gauge, 
and  the  cut-off  0.3.  The  final  pressure  of  the  delivery  air  is  80 
Ibs.  gauge.  The  mean  effective  pressure  for  the  entire  card, 
including  the  period  of  d  scha  ge  represented  by  the  top  horizon- 
tal line,  may  be  obtained  in  columns  7  and  6  respectively.  These 
mean  effective  pressures  represent  the  resistance  to  compression 
offered  by  the  air  under  the  conditions  assumed.  Thus,  for 
example,  according  to  the  table  the  mean  effective  pressure  of  the 
entire  air  card,  assuming  isothermal  compression  to  be  94.7  Ibs. 
pressure  absolute,  80  Ibs.  gauge,  would  be  27.33  Ibs.  per  square 
inch;  if  carried  to  the  same  degree  adiabatically,  36.6  Ibs.  per 
square  inch.  During  the  compression  of  the  strokes  only,  the 
mean  effective  pressures  are  respectively  20.5  and  17.7  Ibs.  per 
square  inch. 


3M  MANUAL  OF  MINING. 

The  Work  of  Adiabatic  Compression  of  w  pounds  of  air  is 

Fa=i83.45  (TI-  rJw-wwp^-PV). 
The  Work  of  Isothermal  Compression  of  w  pounds  of  air  is 

V+v 
Wt=P(V+v)  hyp  log  y~~^-  (P|-P)*w, 

in  which  v  is  the  volume  of  clearance  in  the  cylinder. 

TABLE  OF  HYPERBOLIC  LOGARITHMS. 

(Base  2.72.) 
Hyp  log=2.3O26X common  logarithm. 


Number. 

Logarithm. 

Number. 

Logarithm. 

Number. 

Logarithm 

.01 

.009 

.10 

•74i 

4-60 

1.526 

.02 

.019 

.20 

.788 

4.70 

•547 

•03 

.029 

•3° 

•832 

4.80 

•568 

.04 
•05 

•°39 
.048 

.40 
•5° 

.875 
.916 

4.90 
5.00 

•589 
.609 

.06 

•  058 

.60 

•955 

5.10 

.629 

.07 

.067 

.70 

•993 

5.20 

.648 

.08 

.076 

.80 

.029 

5-3° 

.667 

.09 

.086 

.90 

.064 

5-40 

.686 

.10 

•°95 

3-oo 

.098 

5-50 

.704 

.11 

.104 

3.10 

•131 

5-6o 

.722 

.12 

•"3 

3-20 

.163 

5-7° 

.740 

•13 

.112 

3-3° 

•193 

5.80 

•757 

.14 

•131 

3-40 

•223 

5-90 

•774 

•15 

•139 

3-5° 

•  252 

o.oo 

.791 

.20 

.182 

3-6o 

.280 

6.20 

.824 

•25 

.223 

3-7° 

-308 

6.40 

.856 

•3° 

.262 

3-8o 

•335 

6.60 

.887 

.40 

.336 

3-9° 

.360 

6.80 

.916 

£ 

.405 
.470 

4.00 
4.10 

•  386 
.410 

7.00 
7.20 

•945 
•974 

£ 

•SB? 

4.20 
4-3° 

•435 
.458 

7.40 
7.60 

.001 

.028 

.90 

.641 

4.40 

.481 

7.80 

•°54 

.00 

•693 

4-50 

•504 

8.00 

.079 

The  following  table  will  facilitate  calculation  of  engine-power 
required  for  compression.  These  values  are  to  be  divided  by 
m,  the  modulus  of  the  compressor,  which  is  for  0.50  to  0.70,  to 


THE  COMPRESSION  OF  AIR. 


315 


determine  the  steam-power  required  to  drive  the  piston  against 
friction. 

TABLE  SHOWING  THE  HORSE-POWER  REQUIRED  TO  COMPRESS  AND  DELIVER 
ONE  CUBIC  FOOT  or  FREE  AIR  PER  MINUTE  TO  VARIOUS  GAUGE  PRESSURES; 
ALSO  THE  POWER  REQUIRED  TO  COMPRESS  AND  DELIVER  ONE  CUBIC  FOOT  OF 
AIR  AT  THE  GIVEN  PRESSURE. 


Compressing  One  Cubic  Foot  of 
Free  Air  per  Minute  to 
Given  Pressure. 

Delivering  One  Cubic  Foot  per  Minute 
of  Air  Compressed  to  the 
Pressure  Given. 

I 

Gauge 

Pressure. 

2 

3 

4 

S 

Compression  at 
Constant 

Compression 
without 

Compression  at 
Constant 

Compression 
without 

Temperature 

Cooling. 

Temperature. 

Cooling. 

5 

.01876 

.01963 

.02514 

.0263 

10 

•03325 

.03609 

.05586 

.06399 

IS 

.04507 

.05022 

.09105 

.10145 

20 

.05506 

.06283 

.12994 

.14829 

25 

.06366 

.07422 

.17191 

.20043 

30 

.0713 

.08464 

.21678 

•25734 

35 

.0782 

.09425 

.26445 

.31872 

40 

.084305 

.10324 

•31375 

.38422 

45 

.08954 

,IIl66 

.36368 

•45353 

50 

.09508 

.11952 

.41848 

•52605 

55 

.09936 

.12702 

.47H2 

.60227 

60 

.  10402 

.13418 

•52855 

.68181 

65 

.10808 

.14028 

.58612 

.76079 

70 

.11245 

.14718 

.64812 

.8483 

75 

.11629 

•15373 

.70952 

•93795 

80 

.11926 

•I597I 

.76843 

1.02906 

85 

.1224 

•l6555 

•83039 

1.1231 

90 

.  12558 

.17096 

.89444 

1.2176 

95 

.12886 

.17629 

.96164 

1.3148 

100 

.13121 

•18153 

1.0243. 

1.4171 

EXAMPLES. — i.  Required  the  number  of  units  of  work  necessary  to  compress 
80  Ibs.  of  free  air  from  14.7  Ibs.  absolute  pressure  and  60°  F.  to  a  pressure 
of  88.2  Ibs.  per  square  inch.  Here  P=  2116.8;  7=80X1308=1046; 
Fj=o.i66X  1046=  174.3. 

Neglecting  friction  and  the  value  for  the  clearance  in  the  cylinder, 
JF,=  21 16.8X1046X1. 79=  3,963,320  ft.-lbs.  Compressed  adiabatically,  T=» 
60  +  461  =  521°,  and  TI}  according  to  the  previous  table,  is  nearly  881°,  or  by 
computation  log  t^log  521+0.29  log  (88.2  X 144) —0.29  log  (2116.8)  = 
294254.  Whence  ^=876.1°,  T7a=i83.5(T1-T)x8o=5,2ii,5oo  ft.-lbs. 
By  the  preceding  table  the  work  is  respectively  121.65  and  160.77  h-P-  f°r  * 
final  compression  to  89.7  Ibs.  absolute. 

2.  100  cubic  feet  of  air  per  minute  are  to  be  compressed  to  50  Ibs.  gauge 


316  MANUAL  OF  MINING 

without  cooling.  Required  the  work  expended  in  the  operation.  The  initial 
temperature  of  the  free  air  is  60°;  then  ^=14.7;  P1=64-7;  F=ioo-, 
7i=35-°°;  T=52i°;  ^-339°  F. 

Neglecting  clearance,  the  value  for  the  work  per  cubic  foot,  according  to 
the  table,  is  0.11952  h.p.,  and  for  100  cubic  feet  11.952  h.p. 

100 
Wa=  183.45(800-521)^-^=  11.9  X33,ooo  ft.-lbs. 

If  the  compression  were  conducted  with  perfect  cooling,  then,  according 
to  the  table,  there  would  be  required  9.5  horse-power  and  the  final  volume 
V,  would  be  22.7. 

Wi=  2116.8X100  hyp  log  4.401  =  9.504X33,000  ft.-lbs. 

3.  What  should  be  the  size  of  the  cylinders  to  compress  100  Ibs.  free 
air  per  minute  from  60°  F.  to  100  Ibs.  gauge  pressure?  r=52i°  abs.; 
PI=  114.7X144;  7=1308  cubic  feet;  v=^  per  cent  V;  w=ioo;  ^=946° F. 

PPa=  183.45(946-521)100=  7,797,000  ft.-lbs.; 
Wi=  2116.8(1.07X1308)  hyp  log  ^^-(14,400)91.5=4,843,300  ft.-lbs. 

If  the  modulus,  m,  be  taken  at  0.70,  the  work  to  be  supplied  the  steam- 
piston  is  11,138,571  ft.-lbs.  and  6,919,000  ft.-lbs.  respectively.  With  an 
actual  capacity  of  cy Under  0.90  that  of  the  apparent  capacity,  there  being 
80  strokes  per  minute,  the  cylinder  capacity  is,  for  isothermal  compression, 
16.35  cubic  feet.  With  7  per  cent  clearance  the  cushioned  air  will  correspond 
to  0.3027  of  free  air.  Each  piston  displacement  is  then  (1.302—0.07)8.175  = 
10.07  cubic  feet.  The  cylinders  are  23f"X36"  stroke  each. 

The  mean  effective  resistance  is  30.07  Ibs.  per  square  inch.  The  steam- 
cylinder  must  be  i2$"X36"  if  the  average  cut-off  is  §,  initial  pressure  120  Ibs. 
gauge,  and  the  back  pressure  5  Ibs.  Then,  Chapter  V,  3  per  cent  clearance, 

m.e.p.=o.8658(i2o+ 14. 7)— 5=116.6; 

g=  6,919,000 

33,000 
whence 

jfe=  12.54  inches    and    5=  3  feet. 

The  Mean  Resistance  to  Compression  of  dry  air  during  com- 
pression can  be  ascertained.  Let  Pl  be  the  final  pressure  and 
P  the  initial  pressure;  then 


THE  COMPRESSION  OF  AIR. 


3*7 


Fig.  123  is  a  diagram  showing  on  the  right  the  temperatures 
attained  by  air  when  compressed  adiabatically  to  the  degree 
indicated  at  the  bottom  of  the  figure,  and  on  the  left  the  work 
in  foot-pounds  expended  during  compression.  The  difference 
in  power  consumption  by  adiabatic  compression  and  isothermal 
compression  is  pictorially  revealed.  Thus  a  point  on  the  iso- 
thermal compression  line  corresponding  to  6  atmospheres  indi- 
cates the  work  expended  on  one  pound  as  50,000  ft.-lbs.;  upon 


7          8 

PRESSURES  IN  ATMOSPHERES 


nder 


FIG.  123. — Diagram  Showing  the  Work  Done  During  Air-compression  u 
Various  Conditions. 


one  pound  at  the  same  pressure  adiabatic  there   has  been  ex 
pended  73,300  ft.-lbs.  and  the  temperature  attained  is  460°  F. 

Cooling  the  Air  During  Compression. — Isothermal  compression 
is  more  desirable  to  attain  than  adiabatic  compression,  not  only 
because  the  power  required  is  less,  but  also  because  the  high 
temperature  of  the  latter  is  particularly  injurious  to  the  machinery. 
It  is  an  obstacle  to  rapid  running,  as  proper  lubrication  cannot 
be  maintained. 

Isothermal  compression  is  too  slow.  Moreover,  the  cylinder 
walls  are  never  perfect  conductors.  So  the  air  is  taken  as  cool 


MANUAL  OF  MINING. 

CO 


w    o 


THE  COMPRESSION  OF  AIR.  319 

as  possible — chilled  if  may  be — and  kept  cool  during  all  stages 
of  the  compression  by  artificial  means.  As  the  air  will  cool 
rapidly  after  compression,  producing  a  great  loss  by  radiation 
from  cylinders,  receivers,  conducting  pipes,  and  all  intermediate 
appliances  before  reaching  the  motor,  and  as  it  is  impossible  to 
retain  this  heat  in  spite  of  every  precaution,  it  is  desirable  to 
extract  the  excessive  heat  as  promptly  as  possible.  So  the  supply 
of  air  is  taken  from  outdoors  at  the  coolest  side  of  the  building, 
effecting  a  saving  of  at  least  2  per  cent  in  power  by  this  choice 
of  site,  which  may  be  10°  cooler  than  on  the  other  side.  It  may 
be  noted,  too,  in  the  table  and  the  diagram,  Fig.  123,  that  the 
increment  of  heat  is  greater  in  the  earlier  portion  of  the  stroke. 
During  the  first  15  Ibs.  increase  the  temperature  rises  118°  F.,  but 
during  the  increment  from  90  to  105  Ibs.  the  rise  is  only  37°  F. 

Water-cooling. — Cooling  is  effected  by  a  circulating  current 
of  water  in  order  to  maintain  an  isothermal  compression,  in 
which  case  there  will  be  extracted  from  the  air  under  com- 
pression an  amount  of  heat  corresponding  to  the  area  of  the 
space  between  the  two  curves  in  Fig.  123,  or  an  amount  of  work 
represented  by  the  difference  in  the  vertical  lines  between  the 
isothermal  and  adiabatic  curves  in  diagram  Fig.  122.  The 
cooling  is  a  direct  waste,  and  heat  once  removed  from  the  air  is 
never  returned  to  it.  The  absorption  of  heat  is  accomplished 
by  cold  air  circulating  through  a  jacket,  ;',  Fig.  128,  surrounding 
the  cylinder.  In  Fig.  124,  the  course  of  the  water  through  it  is 
represented  by  the  arrows  and  passages  numbered  i,  2,  3,  and  4. 

A  fine  spray  of  water  injected  into  the  cylinder  would  be  a 
more  effective  method  of  cooling,  were  it  not  so  objectionable 
because  of  its  corrosive  action  on  the  cylinder  walls  and  its  freez- 
ing when,  at  some  later  stage,  it  is  used  expansively  in  the  motor. 
The  spray-injectors  are  more  efficient  coolers  than  water-jackets, 
as  shown  in  the  diagram  Fig.  125,  the  former  curves  being  lower 
than  the  latter. 

The  circulating  water  should  be  the  coolest  attainable,  and 
the  cylinder  lining  as  perfect  a  conductor  as  possible,  to  extract 
the  appreciable  amount  of  heat. 


320  MANUAL  OF  MINING. 

Two-stage  Compression.— The  very  short  time  of  contact 
between  the  water  and  the  hot  air  may  be  increased  by  dividing 
the  air- compression  between  two  cylinders  (Fig.  126),  and  in 
addition  passing  it  through  the  intercooler.  This  latter  is  very 
efficient,  saving  as  much  as  10  per  cent  of  the  power,  which  almost 
counterbalances  the  friction  of  the  machine.  This  compound, 
or  two-stage,  compression  gives  nearly  treble  the  time  of  con- 
tact for  the  circulating  waters.  In  the  larger  cylinder  the  air 
is  compressed  to  one  third  or  one  fourth  its  volume,  and  in  the 


FIG.  125. — Combined  Indicator  Diagrams  from  a  Two-stage  Air-compressor. 

small  cylinder  to  that  required.  The  intercooler  is  essentially 
a  cylinder  with  small  brass  pipes  through  which  water  circulates 
and  around  which  air  flows  from  the  low-  to  the  high-pressure 
cylinder.  In  this  way  a  stepped  curve,  diagram  Fig.  125,  is 
obtained  indicating  economical  results  approaching  an  isothermal 
compression.  From  i£  to  3  horse-power  are  saved  over  the 
single-stage  compression  for  each  100  cubic  feet  of  free  air  at 
the  ordinary  degrees  of  compression  used  in  mining.  Two- 
stage  compression  is  not  advisable  below  4  atmospheres  final 
pressure,  but  is  imperative  for  high  degrees  of  compression. 
But  the  greater  the  degree  of  compression,  the  greater  is  the 
amount  of  abstracted  heat  and  the  more  serious  •  becomes  the 
loss  of  power.  For  i,  2,  3,  4,  5,  and  6  atmospheres,  gauge  reading, 
the  losses  are  28%,  37%,  46%,  50%,  53%,  and  56%,  respectively, 


THE  COMPRESSION   OF  AIR. 


32I 


of  the  original  power.     Economic  work  is  best  obtained  by  operat- 
ing at  as  low  pressure  as  is  consistent  with  the  work. 

HORSE-POWER  NECESSARY  TO  COMPRESS  100  CUBIC  FEET  OF  FREE  AIR  TO 

VARIOUS  PRESSURES  AND  WITH  Two-,  THREE-,  AND  FOUR-STAGE 

COMPRESSORS. 


Horse-power  Necessary. 

Horse-power  Necessary. 

Gauge 

Pressure. 

Pressure. 

Two- 

Three- 

Four- 

Two- 

Three- 

Four- 

stage. 

stage. 

stage. 

stage. 

stage. 

stage. 

100 

15-7 

15-2 

14.2 

900 

36.3 

33-7 

31.0 

200 

21.2 

20.3 

18.8 

1000 

37-8 

34-9 

31.8 

300 

24-5 

23.1 

21.8 

I2OO 

39-7 

36.5 

33-4 

400 

27.7 

25-9 

24.0 

I4OO 

41-3 

37-9 

34-5 

500 

29.4 

27.7 

25-9 

1600 

43-o 

39-4 

35-6 

600 

31.6 

29-5 

27.4 

1800 

44-3 

40-5 

36.7 

700 

33-4 

31.2 

28.9 

2  COO 

45-4 

41.6 

37-8 

800 

34-9 

32.5 

30.1 

2500 

43-o 

39-o 

The  Work  of  Compressing  Moist  Air  is  less  than  that  re- 
quired during  the  comp.ression  of  dry  air,  nor  do  the  mean  net 
resistance  and  the  temperature  increase  as  rapidly. 
RISE  IN  TEMPERATURE  AND  WORK  OF  COMPRESSION  OF  DRY  AND  MOIST  AIR. 


Temperature. 

Work  on  One  Pound. 

Absolute 

Pressure. 

Dry. 

Moist. 

Dry. 

Moist. 

14.7    ' 

68°  F. 

68°  F. 





22 

133-8 

94 

13.300 

13,200 

29.4 

185.9 

in 

23>5°° 

22,500 

36-7 

229.5 

124 

30,500 

29,000 

44.1 

266.7 

135 

37,000 

35,ooo 

51.4 

3°P.2 

145 

43,200 

40,600 

<S.8 

330-1 

153 

48,500 

45,ooo 

73-5 
88.2 

383-5 
428.9 

167 
179 

58,500 
67,160 

52,500 
60,000 

The  mean   effective   resistance  or  pressure  during  compres- 
sion or  expansion  of  moist  air  is 


and 


Kp  \  0.166  ) 

P7       ~x? 


MANUAL  OF  MINING. 


THE  COMPRESSION   OF  AIR. 


323 


The  Air-compressor. — The  simple  principle  of  air-compres- 
sion is  difficult  of  execution.  To  obtain  a  compact  high-speed, 
uniform,  rapid- cooling,  efficient  compressing  engine  is  not  easy. 


The  essentials  are  partly  secured  in  various  ways  by  the  several 
successful  patterns  now  on  the  market. 

The  piston  of  ihe  air- compressor  may  be  driven  directly  by 
a  steam- cylinder  at  the  other  end  of  the  same  rod  (in  tandem), 


324 


MANUAL  OF  MINING. 


or  "  straight  line,"  or  it  may  be  alongside  of  and  joined  by  a 
common  cross-head  with  the  steam-piston.  Such  an  engine  is 
called  a  "  cross"  air-compressor.  It  may  be  driven  by  a  belt  or 
by  a  reducing- gear  from  a  shaft  actuated  by  electricity  or  steam- 
power,  or  an  impulse-wheel  mounted  directly  upon  the  main 
shaft  of  the  compressor.  The  cylinders  are  usually  horizontal  and 


Flo.  128. — A  Water-jacketed  Air-compressor,  IngersoU  Pattern. 

may  be  single  or  duplex,  and  the  air  may  be  compressed  in  one 
stage  with  one  cylinder,  or  in  two  stages,  corresponding  to  simple 
and  compound  steam-engines.  The  duplex  engines  are  simple, 
inhaling  and  compressing  separate  volumes  of  air.  They  have 
none  of  the  merits  of  the  compound  or  two-stage  engines,  in  which 
the  large  low-pressure  cylinder  expels  its  air  into  the  small 
high-pressure  cylinder.  The  cranks  of  a  duplex  compressor  are 
at  quarters  to  equalize  the  rotary  effort. 

The  frame,  compared  with  that  of  the  steam-engine,  is  very 
solid  and  on  heavy  base-plate  Compressors  may  be  had  sectional 
for  convenience  in  transportation.  They  are  fitted  with  auto- 
matic throttling  governors  to  alter  the  steam  pressure  according 
to  the  demand  for  power,  and  are  so  connected  that  when  the 
delivery  air  attains  an  excessive  pressure  it  will  operate  on  the 
piston,  which  raises  ihe  regulator  and  closes  the  governor- valve, 
thus  reducing  compressor  speed.  Their  action  is  very  much  like 
that  of  the  automatic  damper  regulator  mentioned  in  Chapter  IV. 


THE  COMPRESSION  OF  AIR. 


325 


326  MANUAL  OF  MINING. 

High  speed  to  the  piston  is  advantageous  for  the  economy  of 
steam  and  for  capacity,  but  because  of  the  rapid  wear  and  the 
difficulties  with  large  valve  areas,  as  well  as  the  inordinate  resist- 
ance developed  thereby,  the  velocity  is  less  than  300  feet  per 
minute,  except  in  the  larger  sizes. 

The  Burleigh  is  upright,  its  air-cylinder  is  single-acting,  and 
its  peculiarity  lies  in  the  admission  of  steam  one  eighth  of  a  stroke 
before  the  air.  The  Waring  has  a  bonnet,  or  conical  valve,  like 
that  in  Fig.  127,  whose  pistons  are  moved  by  a  rocker  on  the  fly- 
wheel shaft,  the  steam- cylinder  being  set  at  an  angle  to  the  hori- 
zontal air-cylinder.  The  Clayton  has  the  usual  poppet-valve, 
and  is  a  compact  machine,  with  its  fly-wheel  centrally  located. 
The  Delamater  has  an  important  contrivance  for  dropping  the 
discharge-valve  from  its  seat.  This  form  is  very  heavy.  The 
Sullivan,  Ingersoll,  Norwalk,  and  Rand  are  the  popular  pneu- 
matic machines. 

The  Valves  of  the  Compressor  are  of  the  poppet,  spindle,  or 
ring  pattern.  Whatever  their  form,  they  should  open  quickly, 
have  a  full  lift,  and  be  ample  in  size.  Large  inlet-valves  offer 
little  difficulty,  though  for  a  short  time  they  are  subject  to  full 
reservoir  pressure.  An  unrestricted  entry  for  the  air  is  obtained 
easily  by  the  use  of  poppets  held  by  springs.  The  Ingersoll- 
Sergeant  compressor  admits  air  through  a  hollow  piston  and 
rod  (Fig.  128).  This  provides  a  very  liberal  inlet  area,  G,  en- 
ables the  cylinder-covers,  J,  to  be  completely  water- jacketed,  and 
leaves  more  room  for  the  discharge-valves,  H.  The  Norwalk 
employs  a  rotary  valve.  The  inlet-valves  of  the  Rand  are  shown 
at  £,  g' ,  Fig.  127.  They  are  provided  with  guards  that  prevent 
their  falling  into  the  cylinder. 

The  valves  should  be  positive,  and  this  the  poppet  attains, 
though  the  tendency  to  "chattering"  is  the  serious  objection  to 
it,  particularly  for  discharge-valves.  This  arises  from  the  two 
opposing  effort — one,  of  the  air,  to  open,  and  the  other,  of  the 
spring,  to  close  the  valve.  The  valve-gear  shown  in  Fig.  124 
does  away  with  this  trouble  in  the  high-pressure  engines;  the 
arms,  a,  b,  relax  the  spring  pressure  and  allow  of  the  valve  rising 


THE  COMPRESSION   OF  AIR.  327 

full-lift  without  dancing.  Poppet-valves  can  hardly  be  im- 
proved upon  for  low  pressures,  though  their  springs  in  time  lose 
elasticity  and  open  too  soon.  This  reduces  their  efficiency,  as 
also  does  any  slip  of  the  valves.  In  the  Norwalk  pattern  (Fig. 
129)  a  positive  discharge  is  obtained  by  moving  the  valve  by 
cams,  such  that  it  remains  at  rest  till  the  pressure  is  sufficient  to 
open  it  quickly.  A  difficulty  about  this,  it  would  seem,  is  that, 
as  the  reservoir  pressure  constantly  varies  (unless  perfectly  regu- 
lated), the  valves  must  receive  constant  attention.  This  may 
be  corrected  by  an  automatic  governor  like  the  Corliss  release, 
which  will  open  the  valves  at  different  points  in  the  stroke  as 
desired. 

The  discharge-valves  require  careful  construction,  for  their 
leakage  is  equal  to  a  large  clearance  space.  They  are  made 
large,  and,  to  prevent  inordinate  frictional  loss  and  wear,  are  as 
numerous  as  possible.  An  excess  of  air  pressure  over  the  re- 
ceiver pressure  is  necessary  to  open  the  valves  and  expel  the  air. 
This  unavoidable  loss  has  an  important  bearing  upon  the  uni- 
formity of  speed. 

The  Effect  of  the  Clearance  Space. — The  large  clearance 
space  between  the  piston  and  the  cylinder-head  at  the  end  of 
the  stroke  constitutes  one  inevitable  source  of  reduction  of  cylin- 
der capacity.  Not  only  is  the  compressed  air  filling  that  space 
never  discharged,  but  on  the  forward-stroke  it  will  expand  and 
fill  a  volume  that  should  have  been  occupied  by  fresh  atmos- 
pheric air  which  is  being  inhaled. 

If  the  air  in  the  clearance  is  compressed  on  the  return-stroke 
to  60  Ibs.  and  occupies  7  per  cent  of  the  cylinder  volume,  it  is 
evident  that  in  the  forward-stroke  it  expands  to  five  times  the 
original  volume,  or  nearly  one-third  that  of  the  cylinder,  before  any 
fresh  cool  air  whatever  can  be  admitted.  The  effective  volume 
of  the  cylinder  is  therefore  materially  reduced.  Moreover,  this 
air  is  hot,  and,  the  cylinder  also  being  warm,  such  inlet  air  as  does 
enter  is  expanded  to  correspondingly  reduce  the  capacity  of  the 
cylinder.  The  clearance  space  cannot  be  made  smaller  than  TV 
of  an  inch,  and  is  usually  much  more.  The  only  remedy  for  this 


328  MANUAL  OF  MINING. 

loss  is  an  increased  length  of  stroke  or  compounding  the  cylin- 
ders. At  75  Ibs.  absolute  pressure  a  single  cylinder  must  be 
three  times  as  long  as  a  compound  having  the  same  clearance 
loss. 

The  friction  of  the  air-piston  taken  with  that  of  the  valves 
is  about  equal  to  10  per  cenc  of  the  work  of  the  engine,  and 
may  icach  25  per  cent  when  including  the  losses  in  the  steam- 
engine. 

The  Horse-power  of  the  Compressor. — This  is  determined, 
in  the  same  general  manner  as  for  the  steam-engine,  by  the  indi- 
cator cards  or  by  assuming  a  mean  effective  pressure,  obtained 
from  column  6  of  the  table  of  volumes,  and  substituting  the 
same  in  the  general  expression  for  horse-power.  The  mechanical 
efficiency  of  the  air-compressor  m  must  be  taken  at  not  more 
than  60  per  cent  for  the  one- stage  and  70  per  cent  for  the  two- 
stage  cylinder,  measured  in  the  terms  of  the  energy  possessed  by 
the  compressed  air  delivered  from  the  machine,  compared  with 
the  steam-motor  power. 

By  inspection  of  the  diagram  Fig.  122  it  is  seen  that  the 
air  resistance  is  atmospheric  at  the  beginning  of  the  stroke  when 
the  steam-power  is  at  the  maximum.  Toward  the  end  of  the 
stroke  it  is  at  a  maximum,  while  the  expanded  steam  is  at  its  mini- 
mum pressure.  This  large  engine  excess  at  the  beginning,  and 
resistance  at  the  end,  necessitate  the  use  of  heavy  fly-wheels. 
Indeed,  the  fluctuating  stresses  to  which  the  entire  compressor 
is  subject  requires  it  to  be  built  excessively  heavy.  In  the  two- 
stage  compressor  this  inequality  of  engine  excess  is  not  so  marked. 

The  Receiver. — A  receiver  is  a  necessity  in  all  compressed- 
air  systems.  It  is  a  huge  tank  with  regulating  devices  to  main- 
tain uniform  pressure.  In  a  measure  it  is  a  power  accumulator, 
if  of  suitable  size ;  it  compensates  for  the  pulsating  effect  of  each 
stroke  of  the  compressor,  and  for  this  purpose  should  be  within 
50  feet  of  the  compressor.  Another  ought  to  be  placed  near  to 
the  machine  drills  to  reduce  friction  losses.  This  will  serve  also 
as  a  drain,  if  water  is  present,  to  catch  the  water  condensing  in 
the  pipes. 


THE  COMPRESSION   OF  AIR.  329 

Further  cooling  ensues  as  the  compressed  air  is  stored  in 
receiver  or  pipes,  but  represents  a  total  loss  of  power  for  which 
there  is  no  compensation.  A  pound  of .  air  not  cooled  in  the 
compressor,  when  at  80  Ibs.  absolute,  would  radiate  81  B.T.U. 
in  cooling  to  its  initial  atmospheric  temperature  of  60°  F.,  and 
thus  dissipate  63,000  ft.-lbs.  of  work.  Perfect  cooling  in  the 
cylinder  would  have  saved  this  amount  of  steam-power  to  the 
motor.  100  cubic  feet  per  minute  would  thus  have  saved  16 
horse-pcwer,  350  Ibs.  of  steam,  and  45  Ibs.  of  coal  per  hour. 

Transmission  of  Compressed  Air. — The  air  is  conveyed  to 
the  drill,  coal-cutters,  hoist,  etc.,  by  pipes.  With  the  exception 
of  electricity,  no  other  means  of  power  transmission  can  compare 
in  efficiency  with  compressed  air.  The  diameter  of  the  pipe 
is  a  matter  of  the  first  importance. 

The  transmission  losses  appear  in  two  ways :  as  loss  of  power 
and  as  loss  of  pressure,  or  head,  indicated  by  the  difference  in 
gauge  reading  at  the  ends  of  the  line.  There  is  a  distinction 
between  these  two  losses.  The  first  is  the  larger,  due  to  cooling, 
of  the  air  during  compression,  and  is  unavoidable  and  not  charge- 
able to  transmission.  Of  the  power  remaining,  some  of  it  is 
lost  by  subsequent  cooling  and  some  in  overcoming  the  frictional 
resistances  in  the  pipes.  The  power  depends  upon  its  pressure 
and  its  volume.  In  the  process  of  transmission  the  pressure  is 
reduced  by  the  frictional  losses,  but  there  is  a  certain  compensa- 
tion from  a  corresponding  increase  in  the  volume.  The  actual 
loss  of  power  from  this  cause  is,  therefore,  slight  in  ordinary 
mining  conditions. 

Frictional  Resistance  in  Pipes. — The  loss  of  pressure,  or  of 
head,  due  to  frictional  resistances  takes  place  according  to  laws- 
governing  the  flow  of  fluids.  If  the  pipes  be  short,  the  velocity 
will  vary  inversely  as  the  area,  and  the  frictional  loss  will  be 
directly  proportiona"  to  the  square  of  the  velocity  of  the  flow. 
It  will  be  also  proportional  to  the  periphery  and  the  length  of 
the  conduit.  But  in  long  pipes  the  expression  becomes  com- 
plex. Tables  of  the  loss  of  pressure  by  flow  in  pipes  are  given  by 
manufacturers,  and  it  will  be  found  therein  that  air  at  32.8  feet 


33° 


MANUAL  OF  MINING. 


per  second  loses  8.26  Ibs.  pressure  in  a  mile  of  ic-inch  pipe, 
10.04  Ibs.  in  an  8-inch  pipe,  and  20.08  Ibs.  in  a  4-inch  pipe. 

The  table  below,  from  the  handbook  of   the  Norwalk   Iron 
Company,  shows  the  losses  of   pressure  for  given  volumes  and 
velocities  in  pipes  1000  feet  long. 
q  =  volume  of  free  air  passing  per  minute  at  60  Ibs.  gauge  pressure ; 

^  II  II  ((  II  «  II  It    Q_         It  II  « 

p  =  the  loss  of  pressure-head  in  pounds  per  square  inch; 
v  =  the  velocity  in  feet  per  second. 

TABLE  OF  VOLUMES  AND  PRESSURE-HEAD  Loss,  TRANSMITTING  FREE  AIR 
THROUGH  PIPES  1000  FEET  LONG. 


Vel. 

3  Inches. 

4  Inches. 

6  Inches. 

10  Inches. 

V 

P 

9 

<f 

P 

9 

of 

P 

9 

^ 

P 

a 

<t 

J.a8 

.046 

48 

60 

•134 

86 

109 

.023 

193 

244 

.014 

537 

680 

6.56 

.209 

Q6 

121 

•152 

172 

217 

.104 

386 

488 

.064 

i°73 

1359 

0.84 

.488 

144 

182 

.360 

358 

326 

•244 

579 

633 

•145 

1610 

2039 

13.12 

.838 

193 

243 

.628 

343 

436 

.419 

772 

977 

.20 

2146 

2719 

16.40 

I-3I7 

241 

3°4 

.982 

429 

544 

.6S8 

96s 

1221 

•393 

2683 

3399 

19.68 

1.  808 

289 

^64 

i-356 

515 

6S3 

.904 

1158 

1466 

•  542 

3220 

4079 

26.24 
32.80 

3-352 
5.270 

386 
480 

486 
607 

2-513 
3.928 

687 
859 

871 
1088 

1.670 
2-635 

1544 
i93i 

1954 
2443 

1.024 
i-573 

4293 
5367 

5438 
6798 

For  any  degree  of  c.ompression  p",  other  than  60  Ibs.  gauge, 
the  quantity  of  free  air  passing  per  minute  would  be  obtained  by 
the  ratio  74.7  :  ^'+14.^.  This  is  approximately  correct. 


EXAMPLE.  —  Required  the  volume  of  free  air  which  at  70  Ibs.  gauge  pres 
sure  can  be  carried  by  a  6-inch  pipe  with  a  friction  loss  of  only  0.658  Ibs.  per 
square  inch.  From  the  table,  965  cubic  feet  of  free  air  compressed  to  60  Ibs. 
can  be  carried;  hence,  with  air  at  70  Ibs.  pressure,  the  volume  is  965  X  84.  7  -s- 
74.7=1094  cubic  feet  free  air. 

The  Loss  of  Energy  Due  to  Friction.  —  The  measure  of  the 
work  possessed  by  air  is  represented  by  that  expended  in  produc- 
ing the  given  conditions  of  temperature,  pressure,  and  volume. 
Any  loss  in  either  of  these  elements  will  necessarily  diminish  its 
capacity.  During  compression  heat  has  been  abstracted  and 
the  capacity  for  work  is  measured  by  the  amount  of  intrinsic 


THE  COMPRESSION  OF  AIR.  33 1 

energy  remaining.  Its  value  may  be  obtained  from  the  usual 
formulae;  the  pressure,  P,  and  temperature,  T,  at  the  entrance 
to  the  pipe  being  known.  The  volume  occupied  by  one  pound  of 
air  is  then  ascertained: 

P,  ^  =  53.181. 

Then  the  relation  between  the  condition  of  the  air  at  the 
entrance  to  the  pipe  to  that  at  the  point  of  delivery  to  the  motor 
is  ascertained  from  the  formula 

PI/  1.408  _  p  y  1.408 
1^1  — -^2l/2 

If  the  temperature  has  not  changed  during  transmission,  the 
volume  and  pressure  will  bear  the  same  relation  at  the  pipe  exit 
as  at  the  entrance,  and  P2F2  =  53.i8r1.  A  drop  in  the  pressure 
ensues,  however,  due  to  friction  of  flow,  and  this  produces  a 
slight  reduction  in  the  available  energy;  that  is,  P2  is  less  by 
the  frictional  loss  in  the  preceding  table  and  the  volume  F2  has  not 
proportionately  increased,  making  PyVj1'408  less  in  value  than 
that  of  PjTY'408  by  the  amount  of  energy  lost. 

The  friction  in  the  pipe  varies  as  the  velocity  and  the  volume 
of  the  air.  For  a  given  power  the  degree  of  compression  may 
be  increased  in  order  to  reduce  the  volume;  or  the  size  of  the 
pipe  may  be  increased  to  reduce  the  velocity.  The  efficiency 
of  the  system  will  be  increased  by  either  plan.  A  6-inch  pipe, 
for  example,  will  carry  800  cu.  ft.  of  free  air  at  80  Ibs.  pressure 
absolute  a  distance  of  5000  feet  with  a  loss  of  i  Ib.  per  square 
inch.  For  the  same  volume  in  a  4^- inch  pipe  the  loss  of  pressure 
is  5.362  Ibs.  per  square  inch.  These  pipes  would  therefore 
require  receiver  pressures  at  the  pipe  entrance  of  81  and  85.3 
Ibs.,  respectively,  per  square  inch.  The  saving  in  power  with  the 
6-inch  pipe  over  the  4|-inch  pipe  will  therefore  be  3  horse-power. 

In  determining  the  loss  of  head  due  to  given  conditions,  the 
following  formula  is  available: 

n_      V*L 
io,oooD3a' 


MANUAL   OF  MIXIXG. 


in  which  H  =head  or  difference  of  pressure  required  to  overcome 

friction  and  maintain  the  flow  of  the  air; 
V=  volume  of  compressed  air  delivered  in  cubic  feet 

per  minute; 

L  =  length  of  pipe  in  feet; 
D  =nominal  diameter  of  pipe  in  inches; 
a  = coefficient  depending  upon  the  size  of  the  pipe. 
VALUES  FOR  a  AND  Z>5a  FOR  VARIOUS  DIAMETERS  OF  PIPE. 


Nominal 

a 

£»a 

Nominal 

a 

D 

D 

l" 

o-35 

0-35 

3" 

0-73 

177-4 

Ij 

I-525 

0.787 

413-2 

0.662 

5-03 

4 

0.84 

860.2 

2 

4 

0-565 
0.65 

18.08 

1 

0-934 

I.  00 

2918.8 
7776.0 

The  Most  Economical  Size  of  Pipe. — The  friction  in  properly 
designed  pipe  systems  is  not  a  serious  matter  and  can  be  made  as 
small  as  the  most  exacting  requirements  demand,  by  enlarging 
the  pipe  or  securing  a  smooth  interior.  The  capacity  of  a  pipe 
is  somewhat  proportional  to  its  cross-sectional  area,  but  is  affected 
by  the  character  of  its  interior  surface  and  the  various  couplings 
used.  In  calculating  the  size  of  pipe  required,  due  regard  must 
be  paid  to  the  commercial  sizes,  for  their  actual  diameter  is  very 
different  from  their  nominal  diameter.  The  velocity  of  the  air  in 
the  main  pipes  should  not  exceed  25  feet  per  second,  and  the 
dimensions  should  be  determined  accordingly.  At  a  higher  rate 
the  friction  becomes  excessive  and  the  power  lost  in  overcoming 
it  too  large.  Though  the  rate  of  flow  through  the  pipe  is  continually 
increasing  from  beginning  to  end,  all  calculations  for  frictional 
loss  should  bear  this  in  mind,  though  it  may  be  neglected  if  the 
drop  in  pressure  is  small. 

With  due  regard  to  economy  in  installation,  the  size  of  the 
pipe  should  therefore  be  made  as  large  as  advisable  to  reduce 
friction,  but  need  not  be  increased  beyond  that  requisite  to  sup- 
ply a  flow  at  25  feet  per  second.  This  gives,  for  ordinary  mining 
practice,  a  diameter  of  4  inches  for  the  mains  and  not  less  than 


THE  COMPRESSION   OF  AIR. 


333 


2  inches  for  the  stopes  and  rooms.  A  4-inch  pipe  with  air  at 
82-lbs.  gauge  will  supply  five  3-inch  drills  3000  feet  away.  100 
feet  of  i^-inch  pipe  will  serve  for  only  one  drill. 

Compressed-air  Pipes. — The  pipes  used  are  steel-riveted  or 
lap- weld,  as  illustrated  in  Figs.  131  and  132.  The  joints  should 
be  carefully  secured.  Means  must  not  be  neglected  for.  pro- 
viding for  changes  in  length  due  to  alternations  of  temperature. 
Iron  expands  0.000007  its  length  per  i°  F.  This  allowance  is 
more  essential  above  than  below  ground;  and  in  shafts  where 
the  temperature  is  inconstant  compensation-joints  are  neces- 
sary. At  every  300  or  400  feet  a  copper  U  tube  is  attached;  ;is 
flexibility  will  allow  for  contraction  or  expansion  of  the  pipes. 
At  the  Republic  mine  the  brass- lined  expansion- joints  every 
500  feet  allow  for  movements  of  12  inches.  They  rest  on  gas-pipe 
rollers.  The  Chapin  iron-mine  has  expansion-joints  at  every 
680  feet  of  the  24-inch  pipe. 

To  reduce  the  frictional  losses  in  the  transmission  large 
elbows  and  bends  of  long  radii  should  be  used.  The  joints  must 
be  made  very  carefully,  to  reduce  leakage,  unless  the  air-pipes  are 
laid  in  ventilation-passages  of  the  mine.  Leaks  when  discovered 
must  be  plugged.  The  velocity  of  escape  of  compressed  air 
being  20,000  feet  or  over  per  minute,  a  very  large  loss  of  fluid 
will  ensue  from  a  neglected  leak. 

STANDARD   STEAM  AND  EXTRA-STRONG  PIPE  USED  FOR  COMPRESSED-AIR 
HAULAGE  PLANTS. 


Trade 
Diameter, 
Inches. 

Cubic  Feet 
in 
i   Lineal 
Foot. 

Lineal 
Feet 
Necessary 
to  Make 
i  Cubic 

Steam. 

Extra  Strong. 

Actual 
Diameter, 
Inches. 

Thick- 

Weight 

Thick- 

Weight 

Foot. 

ness. 

per  Foot. 

ness. 

per  Foot. 

2 

.0218 

45-41 

•15 

3.61 

.22 

5.02 

2.067 

4 

.0341 

29.32 

.20 

5-74 

.28 

7.67 

2.468 

3 

.0491 

20.36 

.21 

7-54 

•3° 

10.20 

3.067 

3i 

.0668 

15  .00 

.  22 

9.00 

•32 

12.50 

3-548 

4 

.0873 

11.52 

•23 

10.70 

•34 

15.00 

4.026 

4} 

.1105 

9-05 

.24 

12.30 

•35 

I7.6o 

4.508 

5 

.1364 

7-33 

•25 

14.50 

•37 

2O.  50 

5-°45 

Si 

•  165° 

6.06 

.26 

16.40 

.40 

24.50 

5.28 

6 

.1963 

5.10 

.28 

18.80 

•43 

28.60 

6.065 

! 

334  MANUAL  OF  MINING. 

Sleeve  couplings  are  used  in  all  pipe  lines  except  the  wrought 
iron,  which  are  spiral-riveted  or  welded  tubes.  The  joints  are 
carefully  made,  and  leaks  avoided  with  the  greatest  of  care. 
Elbows  of  as  liberal  a  radius  as  possible  must  be  provided  to 
have  the  frictional  resistance  small. 

The  Power  Value  of  Compressed  Air. — In  determining  the 
power  value  of  compressed  air  it  must  be  remembered  that  the 
mean  effective  pressure  of  the  air  is  lower  than  that  of  steam 
for  a  given  cut-off  and  initial  pressure. 

The  mean  effective  pressure  during  expansion  can  be  ascer- 
tained by  substitution  in  the  formula  page  316,  P  being  somewhat 
near  the  atmospheric  pressure.  It  is  less  in  amount  than  is 
the  m.e.p.  of  steam  during  a  similar  expansion,  as  also  is  its 
terminal  pressure.  The  work  it  is  capable  of  equals  the  work 
expended  upon  it  if  the  friction  losses  be  neglected  and  there 
were  no  loss  in  cooling,  for  the  cycles  of  changes  which  the  air 
would  experience  in  expanding  are  duplicates  of  those  during 
compression.  Much  power  has,  however,  been  extracted  from 
the  air,  whose  high  temperature  had  been  reduced  nearly  to  that 
of  the  surrounding  atmosphere. 

The  Work  Performed  while  Expanding  from  a  pressure  P^ 
and  a  temperature  TX,  which  latter  is  near  that  of  the  atmosphere, 
to  a  temperature  r  and  a  pressure  P,  which  is  nearly  15  Ibs.,  is 

JF=  183.45(^1- *)w. 

The  final  temperature,  T,  which  is  far  below  that  of  the  atmos- 
phere, produces  an  intense  refrigeration,  which  is  objectionable 
if  moisture  is  present.  The  moisture  will  be  frozen  in  the  process 
and  may  clog  the  exhaust-passages.  As  this  is  usually  the  case, 
the  air  is  reheated  so  high  that  after  expansion  it  will  still  be 
warmer  than  the  atmosphere.  The  heat  thus  expended  has  the 
effect  of  increasing,  in  proportion  to  the  heat  added,  the  volume, 
and  thereby  the  amount  of  work  obtainable  from  the  air. 

Reheating. — Reheating  may  be  accomplished  by  direct  fire, 
or  by  steam,  passing  through  a  pipe  inside  of  the  air-pipe.  Steam 
thus  used  gives  up  all  of  its  latent  heat,  which  can  be  converted 


THE  COMPRESSION  OF  AIR.  335 

into  an  amount  of  work  in  the  air-engine  far  exceeding  what 
could  be  derived  from  it  if  used  directly  in  the  steam- pump. 

A  reheater-jacket  is  sometimes  used  where  other  means 
are  not  available  or  desirable.  This  method  would  be  preferable 
for  compound  air-engines  which  might  otherwise  require  re- 
heating in  two  stages. 

The  Efficiency  of  Compressed  Air. — As  the  efficiency  of  the 
driven  machine  is  not  over  60  per  cent  and  the  efficiency  of  the 
compressor  is  not  over  75  per  cent,  it  is  evident  that,  with  the 
losses  ensuing  in  pipes,  the  aggregate  efficiency  of  the  combina- 
tion does  not  exceed  25  or  30  per  cent  of  the  original  motor- 
power.  This,  compared  with  the  fuel  burning  at  the  boiler, 
represents  an  efficiency  of  about  2  J  to  3!  per  cent-;  in  other  words, 
the  work  performed  at  the  drill  is  equivalent  to  from  275 
to  385  B.T.U.  per  pound  of  coal.  Notwithstanding  this  waste- 
fulness, compressed  air  serves  well  for  many  purposes  in  mining, 
and  will  retain  its  place  even  against  electricity. 

The  Cummings  System  of  Air  Transmission. — This  is  a  closed 
line  of  pipes  between  the  compressor  and  the  air-engine.  The 
initial  absolute  pressure  of  the  air  entering  the  compressor  and 
the  final  pressure  of  exhaust  from  the  engine  are  about  80  Ibs., 
and  the  terminal  pressure  in  the  compressor  and  the  admission 
pressure  into  the  engine  are  150  Ibs.  This  system  is  used  in 
some  pumping  plants. 

REFERENCES. 

Efficiency  Tests  of  Compressed-air  Machinery,  W.  C.  Popplewell,  Mech. 
Engr.,  Jan.  10,  1903;  Transmission  of  Power  by  Compressed  Air,  W.  C. 
Popplewell,  Mech.  Engr.,  March  5,  1903,  April  18,  1903;  Compressed  Air, 
S.  of  Quart.,  April  1897,  196;  Transmission  of  Compressed  Air  for  Power, 
L.  C.  Bayles,  Compressed  Air,  Nov.  1903;  Hints  Concerning  Air  Compressors, 
H.  H.  Kelley,  Engr.  U.  S.A .,  Mar.  1900;  and  Compressed  Air,  J.  H.  Richards. 

Coal-cutting  Machinery,  E.  W.  Parker,  A.  I.  M.  E.,  Vol.  XXIX,  405; 
The  Lee  Long-wall  Mining  Machine,  H.  P.  Bain,  A.  I.  M.  E.,  Vol.  XXIX, 
474;  Stoping  with  Machine  Drills,  B.  L.  Thane,  A.  I.  M.  E.,  Vol.  XXIX, 
770;  A  New  Type  of  Air  Compressor,  W.  H.  Booth,  M.  &  M.,  Nov.  1903. 


CHAPTER  XI. 

PUMPING. 

The  Water  Seepage  into  Mines. — Water  gains  entrance  into 
mines  by  many  and  devious  ways.  Into  some  workings  it 
flows  incessantly  from  some  watery  stratum,  in  others  the  seepage 
is  intermittent.  .  The  subterranean  current  is  easily  excluded 
from  the  mine  by  the  use  of  a  cement  lining,  or  an  iron  or  steel 
tubing  to  the  shaft,  but  the  seepage  accumulates  and  must  be 
pumped  off,  unless  the  workings  possess  a  natural  drainage  or 
an  easy  effluence  by  adit  or  tunnel  for  the  upper  ground.  A 
gutter  at  the  side  of  the  track,  or  under  the  tramway  path,  with 
a  slope  of  i  in  500,  readily  carries  off  the  water,  and  not  uncom- 
monly delivers  it  to  a  small  wheel  to  drive  a  ventilating-fan. 
Generally  the  seepage,  following  the  hydrodynamic  law,  in- 
creases with  the  depth  of  the  opening,  and  a  very  liberal  sump 
is  provided  for  its  accumulation.  Often  one  shaft  and  its 
workings  become,  naturally,  a  sump  for  the  entire  district,  and 
drain  all  the  neighboring  properties  above  its  level,  and  this 
suggests  a  simple  means  of  keeping  one's  mine  dry.  Otherwise, 
as  the  amount  of  water  to  be  encountered  is  uncertain,  provision 
must  be  made  for  the  handling  of  a  large  volume,  according  to 
the  history  of  similar  properties.  In  some  coal-mines  as  much  as 
4000  gallons  of  water  are  raised  per  ton  of  coal,  and  in  Colorado 
40  tons  of  water  per  ton  of  ore.  The  Chief  of  the  Bureau  of 
Mines  reports  that  during  1902  1721  pumps  delivered  to  the 
surface  water  at  the  rate  of  615,013  gallons  per  minute  from  the 
collieries.  The  magnitude  of  such  work  demands  the  employ- 
ment of  powerful  machinery,  and  often  on  a  plan  too  elaborate 
for  the  means  of  the  average  operator. 

336 


PUMPING.  337 

Methods  of  Unwatering  Mines. — Some  districts  were  drained 
by  a  cooperative  scheme  with  extremely  beneficial  results.  A 
long  tunnel  penetrating  the  country  at  a  level  much  below  the 
lowest  point  of  exploration  drains  considerable  territory,  dis- 
pensing with  the  heavy  individual  plants  and  extending  the  ex- 
ploration and  the  productiveness  of  the  mines.  Numerous  ex- 
amples of  tunnels  miles  in  length  may  be  quoted,  some  even 
carrying  so  much  water  as  to  become  canals  for  transportation. 

Upon  cutting  a  wet  cross-course  to  the  vein,  it  is  a  common 
practice  to  plaster  it  up;  or,  in  encountering  old  workings,  to 
build  a  brick  or  stone  bulkhead,  arched  convex  towards  the 
water  (Fig.  257).  To  provide  means  for  the  escape  of  the  accu- 
mulated water,  which  might  otherwise  do  injury,  a  cast-iron  pipe 
is  built  into  the  dam  near  its  top,  and  another  near  the  bottom. 
Either  or  both  may  be  plugged  as  required.  Similarly,  in 
approaching  abandoned  works,  it  is  required  by  law  in  some 
States  that  a  bore-hole  be  kept  30  to  50  feet  in  advance  of  the 
drift,  with  flank-holes  on  each  side,  to  guard  against  dangers 
from  the  sudden  breaking  into  the  reservoir. 

Under  certain  conditions,  in  stratified  regions,  a  hole  is  drilled 
from  the  sump  down  to  some  permeable  stratum,  into  which 
the  water  is  discharged. 

The  water  entering  a  mine  at  various  levels,  where  economy 
rather  than  simplicity  is  the  object,  should  be  led  to  pumps  at 
the  levels  where  it  issues,  and  not  be  permitted  to  find  its  way 
to  the  bottom,  to  be  raised  the  entire  height  to  the  surface. 

When  the  surroundings  are  such  that  a  tunnel  may  not  be 
used  for  the  unwatering  of  the  mine,  pumping  arrangements 
are  indispensable.  The  earlier  forms  were  crude,  the  engine 
being  of  recent  date.  Surface  waterfalls  were  employed  to 
operate  wheels,  which  raised  bucketfuls  from  below;  or  the 
surface  water  was  arranged  to  compress  air  in  a  reservoir  at  the 
surface,  from  which  pipes  to  the  sump  conveyed  the  compressed 
air,  the  elastic  force  of  which,  in  turn,  forced  the  water  up  to 
the  surface  through  another  pipe.  This  is  a  wasteful  system 
and  intermittent,  but  doubtless  was  cheaper  than  any  other  means 
then  available. 


338 


MANUAL  OF  MINING. 


The  Hydraulic  Ram. — At  the  Comstock  mines  a  special 
hydraulic  ram  is  used,  by  which  1800  gallons  are  pumped  from 
the  26oo-foot  level  to  the  Sutro  tunnel  at  1600  feet.  The  air- 
pressure  in  the  accumulator  is  960  Ibs. 
per  square  inch,  and  the  pipes  at  the 
bottom  sustain  a  pressure  of  2000  Ibs. 

The  efficiency  of  the  ram  diminishes 
with  the  ratio  between  the  quantity  of 
water  raised  and  that  used.  With  a  fall 
of  i  and  a  lift  of  4,  the  efficiency  is  86 
per  cent;  if  the  lift  is  ten  times  the  fall, 
it  is  53  per  cent;  with  i  to  20,  it  is  17 
per  cent;  and  with  i  to  26,  it  is  o. 

Hoisting  Water. — Small  volumes  of 
water  are  handled  by  buckets,  obtainable 
of  any  size,  and  with  a  capacity  up  to 
200  gallons  (Fig.  130).  At  the  bottom  is 
an  inlet-valve  by  which  the  tub  is  quickly 
filled  as  it  sinks  into  the  sump;  it  is  then 
hauled  up,  its  valve  closes,  and  at  the 
surface  it  is  discharged  by  being  brought 
down  on  a  pin  which  again  opens  the 
valve.  In  some  mines  the  water-bucket 
is  attached  underneath  the  cage,  and  travels  continually  with 
it.  Bailing-fanks  holding  500  to  1000  gallons,  balanced  in  pairs 
in  one  compartment,  are  hoisted  by  a  special  drum.  Slopes  are 
equipped  with  similarly  valved  skips,  the  emptying  being  done 
from  the  mouth,  as  with  ore. 

Bailing-tanks  should  always  be  in  readiness  for  immediate 
service  at  every  mine  operated  through  shafts  or  inclines,  to  re- 
lieve the  pump  on  emergency.  Bailing  is  not  economical,  be- 
cause of  the  great  dead  weight  of  tank  and  rope  compared  with 
the  weight  of  water  hoisted.  They  are  adapted  to  a  variable 
inflow  and  give  an  equal,  though  low,  efficiency  with  all  volumes 
of  water. 

Mine-pumps. — The  design  of  mine-pump  plants  requires  the 
greatest  possible  reliability  against  breakdowns,  the  greatest 


FIG.  130. — A  Water-bucket. 


PUMPING.  339 

possible  facility  for  making  repairs,  and,  if  practicable,  the  high- 
est mechanical  efficiency  at  the  normal  rate  of  flow. 

The  pumps  used  in  mining  work  include  reciprocating  pumps 
driven  by  rods ;  steam-,  air-,  or  water-engines ;  rotary  pumps  oper- 
ated by  electricity,  water-  or  steam-turbines,  oil-engines,  or  belts; 
pulsometers,  injectors,  etc.,  employing  steam;  and  siphons. 

The  reciprocating  pumps  include  the  plunger,  piston,  and 
bucket.  Of  these  the  oldest  is  the  lift-bucket,  which,  however, 
affords  no  means  of  repairing  the  barrels  and  packings,  which 
wear  rapidly  with  sandy  waters.  The  piston-pumps  are  suit- 
able for  low  pressures,  but  cannot  be  repaired  as  readily  as  the 
plunger-pumps,  whose  packing  and  stuffing-boxes  are  outside 
and  can  be  replaced  without  stopping  the  pump.  They  may  be 
driven  by  rods  from  a  motor  at  the  surface,  or  by  a  motor  directly 
connected,  receiving  power  through  wires  or  pipes.  The  rod- 
driven  pumps  have  many  disadvantages;  nevertheless,  even 
with  the  competition  from  steam  and  air-driven  motors,  they  are 
much  used. 

Belts  or  hemp  rope  may  also  be  used  to  drive  pumps  on  a 
separate  foundation.  They  are  called  power-driven  pumps, 
having  the  additional  advantage  that  they  may  be  located  remote 
from  the  power  source.  Of  these  the  rotary  pumps  and  centrif- 
ugals have  a  capacity  which  can  be  economically  varied  within 
very  narrow  limits. 

Injectors,  pulsometers,  ejectors,  etc.,  are  to  be  regarded  as 
temporary  substitutes  only  during  the  repairs  to  the  main  plant. 

Mine  Rod-pumps. — The  rod-driven  pumps  are  of  two  types. 
One,  the  lift-pump,  is  a  single  line  of  rods,  which  reciprocates  in 
a  straight  line  of  pipe.  The  lift-pump  is  a  direct-driven  pump, 
with  vertical  single-acting  engine  by  which  the  water  and  the  rod 
are  lifted  on  the  up- stroke,  while  the  weight  of  the  latter  carries 
it  down.  The  other,  the  Cornish  system,  is  an  extended  rod 
operating  outside  of  the  column-pipe  upon  a  number  of  plunger- 
pumps  along  the  line,  with  a  lift  section  at  the  bottom.  In  the 
Cornish  pump  the  water  is  forced  up  on  the  down-stroke  by  the 
weight  of  the  rods.  The  engine  at  the  surface  raises  the  rods 


34°  MANUAL  OF  MINING. 

and  plungers  besides  lifting  the  water  in  the  suction-lift.  The 
Cornish  system  employs  a  rotary  engine,  making  it  possible  to 
utilize  the  expansive  force  of  the  steam. 

Pump-pipes. — The  pipe  through  which  the  water  flows  is 
variously  termed  as  a  stand-pipe,  column-pipe,  and  lift.  It  is 
of  a  diameter  commonly  10  inches,  often  as  much  as  20,  and  ex- 
tends from  bottom  to  top.  It  is  of  cast  iron, 
wrought-iron  lap-weld  steel  or  spiral  riveted, 
or  welded,  in  standard  lengths  of  5  to  20 
feet.  The  cast-iron  pipe,  having  a  smooth 
interior  and  uniform  diameter  throughout,  is 
preferable  and  more  convenient  than  tha 
riveted  pipe  (Fig.  131)  or  the  lap-weld  iron 
(Fig.  132);  but  as  it  represents  too  much 
dead-weight  for  the  strength,  its  days  of 
utility  are  nearing  an  end.  The  ideal  pipe 
is  of  steel,  which  gives  the  lightest,  strongest, 
and  most  durable  tubing;  this  may  be  had 
in  four  grades,  light  to  extra  heavy.  It  is 
made  of  sheet  steel  spirally  laid,  riveted 
at  the  overlapping  joints  or  cold  hammer- 
welded.  The  pipes  are  united  by  bolting 
together  at  the  flanges,  which  are  riveted, 
screwed,  or  locked  on  the  pipe  (Fig.  132);  or, 
preferably,  they  are  coupled  on  the  hub-and- 
spigot  plan  of  sleeve  (Fig.  133).  This  is  a 
double  socket,  into  which  the  pipe  is  slipped, 
"oakumed,"  and  leaded  from  each  side,  as 
shown.  For  this  joint  the  pipes  have  expanded 
ends. 

A  water-tight  joint  is  secured  by  placing 
rubber,  leather,  lead,  or,  best  of  all,  corrugated 
copper  gaskets  between  the  flanges,  which  are 

FIG.  131.— Spiral  Pipe.     .  rr  ,     ,  ,-i     i  •  > 

then  bolted  together  while  lowenng.  Spence  s 
metal,  used  as  a  calker,  offers  an  excellent  joint,  is  cheaper  than 
lead,  and  ought  to  be  better  known. 


PUMPING.  341 


The  pipes  last  from  fifteen  to  twenty  years  unless  the  water  is 
corrosive,  in  which  case  gun-metal  is  used.     If  the  water  is  very 


FlG.  132.— Pipe-joint. 

bad,  wooden  pipes  are  made  by  hollowing  the  trees,  fitting  the 
joints,  tarring  them,  and  strengthening  by  wrought-iron  bands 
at  every  3  to  6  feet.  In  many  mines  recourse  has  been  had  to 


FIG.  133. — Pipe-join.-,. 

these  as  the  only  stand-pipe  that  will  last  over  six  weeks. 

In  calculating  ^ the  flow  of  water  through  pipes,  the  effect  of 
entrance  head,  the  friction  factor,  the  effect  of  bends,  elbows, 
tees,  valves,  gates,  etc.,  as  well  as  the  condition  of  pipe-joints 
and  the  internal  surface  of  the  pipe,  should  enter  into  the  calcu- 
lations; otherwise  the  results  obtained  are  mere  approxima- 
tions. 

Care  should  be  taken  to  give  independent  support  to  the 
water-pipe  lines  to  prevent  motion,  to  relieve  the  lower  sections 
of  the  pressure  from  above,  and  to  avoid  the  evil  consequences 
of  vibration.  They  are  always  stayed  laterally  to  keep  them  in 
line.  In  no  case  should  the  supports  be  rigidly  connected,  pre- 
venting expansion  or  contraction.  All  bends  and  elbows  should 
be  supported,  as  the  curve  of  the  pipe  receives  excessive  pressure 
and  is  also  of  weak  resistance.  All  the  bands  are  made  of  cast- 


342  MANUAL  OF  MINING. 

ings,  and  the  long  bands  may  be  of  riveted  pipe.  The  acu'e- 
angular  turns  are  made  by  inserting  thick  wedge-gaskets. 

The  size  of  the  pipe  should  be  as  large  as  possible,  that  the 
velocity  of  flow  may  be  maintained  with  little  resistance.  The 
suction-pipe  should  be  short  and  more  liberal  in  diameter  than 
the  delivery.  The  velocity  of  water  in  the  pipe  differs  little  from 
the  velocity  of  the  plunger  or  piston.  Water-hammer  and  seri- 
ous consequences  would  be  the  result,  if  water  were  allowed  to  flow 
intermittently  through  pipes  at  more  than  300  feet  per  minute. 

Owing  to  the  corrosive  action  of  most  waters  and  their  solid 
contents,  the  pipes  are  made  as  thick  as  consistent  with  other 
conditions  in  the  problem.  The  varying  pressures,  the  acidity 
of  water,  the  water-hammer,  and  the  weight  of  the  column  which 
is  supported,  are  elements  which  increase  the  thickness  of  the 
pipe. 

Valves  on  Pump-pipes. — Air-chambers  are  supplied  at  con- 
venient points  of  long  hydraulic  pipe,  if  no  provision  is  made  for 
the  insertion  of  relief-valves  or  pressure-regulators.  The  air- 
chambers  should  be  made  air-tight  and  coated  with  heavy  asphalt 
for  heavy  pressures.  Relief- valves  are  spring-loaded  or  weighted 
pistons,  or  valves  set  to  open  at  a  given  pressure.  The  former 
are  more  sensitive  than  the  latter.  Ample  lift  should  be  given 
to  allow  of  a  prompt  discharge  of  air  or  of  water  to  relieve  the 
excess  pressure  promptly. 

Automatic  air-valves  on  hydraulic  pipe  lines  are  sometimes 
used,  constructed  so  that  they  will  close  by  the  combined  effect 
of  buoyancy  and  the  velocity  head  of  the  water.  Some  air- 
valves  are  provided  to  admit  air  and  prevent  a  collapse  of  the 
pipe  from  atmospheric  pressure  when  the  pipe  is  emptied  of 
water.  They  do  not  operate,  however,  to  let  air  into  the  pipe 
until  the  pressure  falls  very  low.  The  pipes  should  be  located 
in  the  up-cast  shaft  compartments,  the  steam  and  reheated  air- 
pipes  being  protected  also  against  radiation.  Steam-pipes  and 
air-pipes  are  provided  with  traps  at  low  points  for  drains. 

Check-valves  must  not  be  neglected  where  the  conditions  are 
such  that  a  stoppage  in  the  pump  may  cause  a  reverse  flow  of 


PUMPING.  343 

the  water  with  consequent  danger  to  machinery.  All  valves  and 
gates  and  water-pipes  should  close  tightly  and  without  shock. 

Pump-valves. — The  valves  for  pumps  are  hinged,  called 
clack-valves,  straight  lift-valves,  rising  vertically  from  their  seats, 
and  flexible  valves,  which  alter  their  form  on  opening.  Direct- 
driven  pumps  employ  straight  lift-valves  of  rubber  or  vulcanite 
resting  on  a  grating.  The  requirements  of  a  good  valve  are  that 
it  should  close  promptly  and  perfectly  on  its  seat,  open  easily, 
-and  remain  with  a  minimum  of  pressure.  It  should  not  present 
much  resistance  to  the  flow  or  divert  the  current  from  a  straight 
line.  It  should  be  simple  and  accessible. 

If  the  waters  carry  material  in  suspension,  the  whole  valve 
should  be  made  of  some  elastic  material  in  order  that  the  par- 
ticles lodging  on  the  seat  shall  not  spoil  the  band.  Leather  can- 
not be  used  if  the  water  is  acid.  Rubber  compositions  are  em- 
ployed for  hot  water.  Hinged  valves  are  more  liable  to  leak 
than  straight  lift-valves.  The  use  of  wooden  seats  for  metal- 
faced  valves  is  objectionable  because  they  leak,  although  they 
are  more  desirable  than  rubber  or  leather.  For  acid  waters  the 
seats  are  usually  of  brass  screwed  into  position,  and  thus  easily 
removed  when  too  far  gone  from  corrosion.  The  straight- line 
lift-valve  (Fig  141)  of  rubber  composition  rising  from  a  grating 
and  held  in  brass  cages  is  used  for  pressures  up  to  500  Ibs.  to  the 
square  inch,  but  above  that  the  discs  are  of  metal  with  leather 
facing.  Flexible  valves  are  generally  of  rubber,  suitable  only 
for  moderate  lifts,  and  round  in  form,  or  rectangular  with  all  of 
the  corners  trimmed. 

The  area  of  the  valves  is  large  to  allow  of  a  free  flow  of  the 
"water,  but  increasing  the  area  of  the  valve  increases  the  pressure 
upon  its  surface  and  requires  a  heavy,  thick  valve.  Hence  it  is 
more  desirable  to  use  a  large  number  of  small  valves  than  to 
employ  few  of  large  area.  On  the  other  hand,  it  is  desirable  to 
restrict  the  lift  of  the  valves  as  much  as  possible  and  to  reduce 
the  velocity  and  consequent  resistance  of  the  flow  past  the  valve 
to  a  minimum.  Enlargement  of  the  area  increases  the  risk  of 
leakage  and  also  the  percentage  of  slip.  Again,  increasing  the 


344  MANUAL  OF  MINING. 

area  of  the  valve  increases  the  width  of  the  bearing  and  its  seat. 
This  makes  the  difference  between  the  lower  area  of  the  valve 
subject  to  pump  pressure  and  the  upper  area  of  the  valve  sub- 
ject to  the  resistance  in  the  pipe  so  great  that  an  excessive  over- 
pressure will  be  required  to  raise  it. 

The  Working-barrel. — At  the  lower  end  of  the  stand-pipe  a 
12-foot  length  of  cast  iron  constitutes  the  working-barrel,  in 
which  oscillates  a  piston  carrying  an  upward-opening  valve, 
similar  to  that  at  the  lower  end  of  the  barrel  (Fig.  136).  For 
acidulous  waters  the  barrel  is  bushed  with  gun-metal.  It 
should  be  thick,  to  admit  of  being  bored  out  several  times,  as  it 
is  rapidly  cut  away  by  the  gritty  waters  during  sinking. 

The  working-barrel  can  never  be  more  than  28  feet  from  the 
sump-level;  in  mountainous  districts  still  less;  at  5600  feet  alti- 
tude, 23  feet;  and  at  10,000  feet,  18  feet.  Usually  the  working- 
barrel  and  suction-pipe  are  suspended  by  chains  from  two  stulls 
resting  in  the  cribbing,  and  the  stand-pipe  supported  at  intervals 
by  stout  reachers. 

The  Suction  Length. — Below  the  barrel  is  a  length  of  pipe 
or  flexible  hose  dipping  into  the  sump  and  receiving  the  water 
through  a  perforated  strainer.  During  sinking  this  suction-pipe 
must  follow  the  lowering  of  the  sump.  During  the  blast  it  is 
raised  for  each  shot  or  boarded  over. 

The  strainer  prevents  sand  and  gravel  entering  the  cylinder. 
The  suction-lift  must  be  as  short  as  possible  and  as  large  as  ad- 
missible. As  the  water-level  lowers  during  the  sinking  opera- 
tions in  a  mine,  a  stationary  pump  should  be  provided  with  a 
wire-wound  flexible  hose  below  its  suction-lift,  dipping  into  the 
sump,  or  a  telescopic  joint  allowing  for  a  supply  of  10  feet.  The 
hose  is  preferable  because  it  can  be  adjusted  to  the  bottom  of 
the  shaft.  Owing  to  injury  and  hard  usage  it  is  usually  covered 
with  canvas  and  wire- wound. 

When  the  water-level  has  receded  below  the  allowable  suc- 
tion height  of  the  pump  a  length  of  pipe  must  be  added  above 
the  pump  which  is  then  lowered  to  a  degree  determined  by 
fixed  conditions.  Under  such  circumstances  the  working-'oump 


.    PUMPING.  345 

is  suspended  by  chains  from  two  stulls  resting  in  the  cribbing, 
the  stand-pipe  above  the  pump  being  independently  supported 
at  intervals. 

The  Lift-pump.— The   piston,  or  "bucket,"  is  attached    by 
an  iron  fork  (Fig.  134)  to  a  wooden  rod,  4  or  5  inches  square,  ex- 


FIG.  134. — Single-acting  Steam  Lift-pump. 

tending  up  through  the  pipe  to  the  surface,  where  it  is  connected 
either  with  one  end  of  a  walking-beam  or  to  the  piston  of  a  single- 
acting  engine.  As  it  receives  a  tensile  strain,  the  joints  are  scarfed 
and  strapped,  or,  if  the  ends  are  flushed,  two  continuous  lines  of 
strap-iron,  breaking  joints,  are  bolted  together  through  the  rod. 
The  latter  plan  reduces  the  breakage  and  the  number  of  stop- 
pages for  repairs.  A  4- inch  rod  is  large  enough  for  a  1 2-inch 


346  MANUAL  OF  MINING. 

pipe,  and  a  5-inchrod,  properly  spliced  and  strapped,  for  a  1 3-inch 
to  i6-inch  deliver}'.  The  size  of  the  straps  is  easily  calculated. 
The  area  of  each  one,  a,  is  o.ooooj$d2D.  A  2oo-foot  pump-rod 
requires  two  straps  4X£  or  3X{  for  a  lo-inch  pipe. 

On  the  down- stroke  the  rod  falls  through  the  column  of 
water,  while  the  valve  in  its  piston  opens  and  the  clack  of  the 
wcrhing-barrel  closes.  Returning,  the  valve's  action  is  reversed, 
water  rises  from  the  sump  into  the  working-barrel,  and  all  that 
above  the  piston  is  lifted  a  distance  equal  to  the  stroke,  and  a 
column  of  water  simultaneously  discharged  at  the  surface. 

At  the  surface  the  column-pipe  terminates  in  an  elbow  dis- 
charge or  in  a  laundry  box  and  trough,  the  pump-rod  continuing 
up  to  the  framing.  The  mechanism  by  which  the  motion  is 
communicated  to  it  is  simple.  A  stout  frame,  with  two  samson 
posts,  supports  a  walking-beam  receiving  its  oscillatory  motion 


FIG.  135. — Connection  for  Lift-pump  Rod. 

from  a  pitman  actuated  by  a  crank-arm  adjustable  to  a  i-,  2-,  or 
3-foot  radius,  giving  strokes  of  double  this  length,  at  the  opposite 
end,  to  the  pump-rod,  which  requires  little  force  besides  its  own 
weight.  The  arm  is  on  a  shaft  turned  from  the  engine  by  cog, 
geared  i  to  6  or  7,  giving  12  to  20  strokes  per  minute  to  the  rod. 
The  ironwork  of  this  frame,  inclusive  of  cogs,  pulley,  and  cast- 
ings, will  cost  about  $250;  the  woodwork,  including  a  24-ft.XT5 
inches  square  walking-beam,  about  $125. 

Lift-pump  Plunger. — The  valves  are  made  of  several  thick- 
nesses of  oak-tanned  leather  cut  into  discs,  tacked  together,  and 
slipping  easily  on  a  grid  at  the  top  of  a  cast-iron  cellular  ring- 
bucket.  A  perforated  cast-iron  guard  on  the  grid  limits  the 
rising  of  the  valve  as  the  water  passes  through  the  bucket.  The 
cellular-ring  bucket  casting  is  all  there  is  of  the  piston,  which 
Slides  freely  in  the  barrel,  and  has  no  other  packing  than 'that 
offered  by  the  leather  discs  forming  the  valve  and  which  are  cut 
larger  than  the  cylinder.  The  rapid  movement,  the  wear,  par- 


PUMPING. 


347 


ticularly  during  sinking,  and  the  heavy 
pressure  upon  these  valves  consume 
a  set  every  two  weeks,  and,  though 
substitutes  have  been  suggested,  such 
as  flexible  brass  or  gutta-percha  plates, 
they  are  not  more  durable,  nor  have 
the  brass  balls  or  conical  poppets  had 
any  marked  success.  The  valves  are 
repaired  or  replaced  by  removing  a 
bolted  door-plate  in  the  barrel  opposite 
them  (Fig.  136). 

The  Direct  Lift-pump.  —  The  frame- 
work just  described  is  crude.  A  simpler 
connection  is  a  vertical  steam-  cylinder 
placed  over  the  shaft  wi  h  its  piston  bolt- 
ed to  a  fork  on  the  rod  This  dispenses 
with  framing  (Fig.  134)  beyond  that  nec- 
essary for  foundation.  The  piston  may 
be  single-acting,  though  frequently  the 
steam  is  admitted  on  both  sides  These 
cylinders  are  easily  set,  their  cost  is  low, 
and  their  maintenance  is  small.  Cylin- 
ders of  i2"X3o"  are  operated  at  a  rate 
of  24  double  strokes  per  minute.  The 
Bull  pump  is  an  example  of  this  type, 
well  known  in  collieries.  It  is  single- 
acting,  and  is  as  large  as  60  inches  in 
diameter,  making  6  to  8  strokes  of  10 
feet  each. 

It  is  not  certain  that  this  form  of 
pump  gives  a  higher  duty  than  the  drive- 
rod  pump.  Except  for  increasing  the 
speed  there  is  no  occasion  for  using  steam 
on  the  down-  stroke.  An  economic  degree 
of  expansion  on  the  up-stroke  is  not  pos- 
sible. The  main  objection  to  its  more 
general  adoption  is  the  large  area  of 


36.—  Drive-rod  Plunger. 


•348  MANUAL  OF  MINING. 

shaft-mouth  it  covers.  Besides,  to  lengthen  or  repair  the  rod  or 
column-pipe  the  cylinder  must  be  displaced,  or  the  additions 
are  made  below;  either  is  slow.  This  pump  cannot  be  used  in 
slopes;  the  irregular  wear  of  the  cylinder  on  one  side  cannot  be 
compensated  for,  nor  the  friction  of  the  rod  in  the  pipe  coun- 
teracted. 

The  Capacity  of  Lift-pumps. — If  d  be  the  internal  diameter 
of  the  pipe  in  inches,  and  L  the  stroke  in  feet,  the  discharge 
in  gallons  with  each  upstroke  is  0.0408 d2L;  and  the  work 
done  per  minute,  in  foot-pounds,  exclus've  of  resistance  in 
the  cog-gear  and  the  mechanism  for  transmitting  the  power, 
is  0.3427  2d2LND,  where  N  is  the  number  of  double  strokes  per 
minute,  and  D  the  height  of  the  water-column  in  feet.  The 
direct-connection  lift-pump  wastes  less  power  in  friction  and 
has  an  efficiency  of  about  85  per  cent.  Its  least  working  steam- 
pressure  is  represented  by  this  equation : 

k2p=o.34$d2D    and    ^-0.551(0.7854^— a)Z>. 

With  a  moderate  steam  pressure  the  ratio  between  k  and  d 
becomes  large  if  the  shaft  is  deep,  and  the  area  of  the  wooden 
rod,  a,  must  be  large  enough  to  support  the  load.  When  the 
depth  of  the  shaft  has  reached  250  feet  the  lift-pump  is  no  longer 
practicable,  and  must  be  altered  to  a  single  discharge-force,  or 
replaced  by  the  more  economic  continuous-flow  steam-pump. 

Force-pumps — Rods. — The  pump-rod  of  the  force-pump,  unlike 
that  of  the  lift -pump,  works  outside,  not  inside,  of  the  stand-pipe. 
Its  lower  end  is  bolted  to  one  end  of  a  cast-iron  H  chamber,  the 
other  stem  of  which  carries  a  long  working-barrel,  into 'which 
plays  a  solid  plunger-piston  (Fig.  136),  instead  of  the  bucket- 
valve  of  the  lift-pump.  Below  the  stem  carrying  the  stand-pipe 
is  the  suction-pipe,  and  in  it  is  an  upward-lift  clack-valve.  Then 
during  the  up-stroke  the  lower  clack  opens,  while  water  rises 
through  it  into  the  working-barrel;  as  the  plunger  falls  the  water 
is  driven  through  the  upper  valve  against  the  column  of  water 
in  the  stand-pipe.  As  sinking  progresses,  suction  lengths  are 


PUMPING.  349> 

added  at  the  bottom  until  the  sump  is  lowered  beyond  the  suc- 
tion distance,  then  the  pump  is  lowered,  while  additional  lengths 
of  pipe  are  attached  between  it  and  the  H  piece  at  the  discharge- 
station  above,  until  a  lift  of  300  feet  has  been  reached;  then  the1 
suction-barrel  is  removed,  to  do  service  similarly  for  the  lower 
lifts,  and  is  replaced  by  another  H  piece. 

The  Cornish  Pump. — The  Cornish  system  is  a  develop- 
ment of  the  rod-driven  pump  by  which  the  lifting  section  is  added 
to  a  series  of  force-pumps  in  a  continuous  line,  the  water  being 
driven  by  stages  to  the  surface  (Fig.  138).  The  engine  raises 
the  rod,  and  water  is  sucked  up  into  the  working-barrel  at  the- 
bottom,  while  that  above  the  bucket  is  lifted  to  the  first  station) 
above.  Here  the  plunger  is  drawing  water  up  through  the  lower 
clack  into  its  barrel,  all  the  other  plungers  doing  the  same.  At 
the  end  of  the  stroke  (6  to  10  feet)  occurs  a  slight  halt,  incidental 
to  the  change  of  direction.  The  rod  falls  by  reason  of  its  own 
weight,  and  each  plunger  closes  the  lower  clack  in  its  H  piece, 
opens  the  upper  one,  and  forces  the  water  out  of  its  working- 
barrel,  driving  at  the  same  time  the  entire  lift-column  an  amount 
equal  to  that  of  the  stroke.  At  the  surface  a  volume  of  water  is 
discharged  on  the  down-stroke. 

The  Cornish  Rod. — The  pump-rod  extends  down  the  shaft, 
and  terminates  at  the  bottom  in  the  piston  of  the  bucket-lift. 
At  intervals  are  offsets,  to  which  are  bolted  the  rods  carrying 
external  plungers,  not  pistons,  to  move  the  water,  and  are  there- 
fore more  easily  packed  and  admit  of  pumping  against  higher 
heads,  and  remaining  tight  much  longer,  than  piston-pumps, 
being  much  less  subject  to  wear.  They  are  made  of  cast  iron, 
though  brass  is  a  better  material,  having  also  less  friction.  The 
stuffing-boxes  for  packing  the  plunger  are  separate  from  and 
built  to  top  of  the  pump-level.  The  usual  packing  is  of  square 
braids  of  hemp,  flax,  or  cotton  soaked  in  tallow  or  a  mixture  of 
tallow  with  beeswax.  Flax  is  the  cheapest,  but  hemp  is  more 
durable.  Both  of  them  offer  greater  friction  than  leather  as  a 
packing.  The  latter  is  used  in  all  pistons  and  bucket-pumps. 

Each  plunger  reciprocates  within  its  working-barrel,  whiclr 


-.35°  MANUAL   OF  MINING. 

is  one  leg  of  an  H  piece  (Fig.  137).  At  every  station  is  located 
;an  H  piece  having  at  the  top  and  the  bottom  a  hinged  valve 
opening  upward.  The  working-barrel  is  about  15  feet  long. 

Excepting  the  short  suction-lift  length  at  the  bottom,  the 
column-pipe  is  in  a  continuous  line,  broLen  only  for  the  intro- 
duction of  H  pieces,  or  tanks,  at  ;he  stations. 

The  lifts  are  rarely  over  300  feet  or  less  than  150  feet  in 
'.height.  As  the  stations  require  heavy  timbering,  especially  if 
•.tanks  are  used,  their  dimensions  must  be  large  and  their  sup- 
ports solid  and  independent  of  the  shaft- timbers.  They  are  not 
any  more  numerous  than  the  circumstances  demand.  The 
greater  the  number  the  greater  the  speed  of  pump  and  the  smaller 
the  diameter  of  the  pipe  will  be  for  a  given  first  cost,  exclusive  of 
that  of  the  tank-chambers. 

The  pump-iod  is  composed  of  sections  joined  at  their  ends 
by  iron  straps  to  bear  the  continual  reversal  of  direction  of  stress. 
Their  length  is  as  great  as  convenient  to  handle,  and  their  sec- 
tion as  large  as  necessary  to  resist  the  tensile  and  compressive 
forces  to  which  they  are  subject.  On  the  up  stroke  all  sections 
are  in  tension.  On  the  down-stroke  the  net  stress  is  the  resultant 
of  the  compression  from  forcing  the  water  and  the  tension  of 
the  pendent  weight  below  the  section,  modified  by  the  inertia  of 
the  attachments  and  counterweights  The  counterbalances 
reduce  the  tension  and  increase  the  compression  by  an  equal 
amount. 

The  aggregate  weight  of  the  rod  is  greater  than  that  of  the 
water  to  be  moved.  Its  cross-section  is  smallest  at  the  bottom, 
increasing  to  the  top  with  the  increase  of  tension  upon  a  section 
from  the  pendent  weight  below. 

To  prevent  accident  from  breakage  or  buckling  of  the  rod, 
stout  stulls  are  laid  across  the  shaft  at  intervals  close  to  the  rod 
to  catch  corresponding  "wings"  of  heavy  timber  clamped  by  iron 
collars  on  the  rod.  The  stulls  are  called  '  guides,"  or  "  stays." 

In  a  deep  mine  requiring  many  superposed  sets  cf  pumps  on 
the  same  rod  the  stroke  of  the  lowest  length  is  shorter  than  that 
of  the  topmost  plunger.  This  may  require  an  increasing  area 


PUMPING  35  r 

of  the  plungers  from  bottom  to  top.     Usually,  however,   they 
are  the  same.     It  not  infrequently  happens  that  there  is  a  differ- 


FlG.  137. — An  H-piece  Cylinder  of  a  Cornish  Pump. 

ent  displacement  of  the  pumps,  which  may  cause  pounding  or 
other  troubles.    A  float  provided  at  each  station  to  tap  the  pipe 


35 2  MANUAL  OF  MIXING. 

into  the  station  tank  when  it  happens  to  be  drained  too  low. 
•would  maintain  a  supply  of  water  and  prevent  overflow  of  one 
tank  and  the  draining  of  another. 

Elastic  bumpers  are  also  needed  to  break  the  force  of  the  blow 
which  might  occur  with  a  stroke  greater  than  the  intended  limit 
through  a  variation  of  the  steam  pressure  or  neglect  in  regulation 
of  the  water  flow.  As  an  instance  of  the  size  of  a  rod — that  of 
4he  Maira,  2300  feet  deep—we  find  the  first  780  feet  down  was 
36"X32",  tapering  to  12" X  24";  at  864  feet  it  was  16  feet  square; 
it  tapered  to  14  inches  square  at  964  feet;  thence  it  was  13  inches 
and  at  the  bottom  12  inches.  In  well- ventilated  shafts  wood  is 
the  preferable  material  for  rods,  neither  wrought-iron  rods  nor 
•wire  rope  having  the  requisite  resilience. 

In  vertical  shafts  the  rods  fall  freely  by  their  own  weight; 
in  slopes  they  rest  on  friction-rollers,  placed  about  30  feet  apart. 
When  iron  ropes  are  used  instead  of  wooden  rods,  sheaves  sup- 
port them.  Changes  in  the  slope  may  be  provided  for  by  the 
use  of  the  rocking-arm.  A  chamber  is  cut  in  the  shaft  at  the 
angle,  in  which  is  firmly  set  a  frame  on  which  swings  a  V  bob 
"by  a  hinge-pin  at  the  apex.  To  the  two  arms  of  the  angle  the 
inclined  rods  are  attached.  While  this  arrangement  is  not  de- 
sirable, because  of  the  expense  and  the  loss  of  power,  still  it  is 
the  best  to  be  had  when  slopes  are  sunk  on  contorted  veins.  The 
application  of  the  Cornish  system  to  inclines  is  attended  with 
many  drawbacks. 

In  mines  utilizing  the  pump-rod  for  a  man-engine  additional 
counterbalance  weights  are  connected  at  intervals  down  the  shaft 
{Fig.  138).  Sometimes  two  lines  of  rods  are  used  in  a  shaft, 
working  two  pumps  from  the  same  bob,  in  which  case  no  counter- 
balance is  needed. 

The  ironwork  of  the  rod  should  be  protected  against  rust, 
particularly  the  joints  and  the  inside  of  the  hollow  iron  pump-rods. 
Pickling  in  acid  will  remove  the  rust,  after  which  a  coating  with 
•warm  oil  and  red  lead  is  recommended. 

The  Balance-bob. — At  the  surface  the  rod  is  directly  con- 
nected to  a  pin  at  one  apex  of  a  king-post  balance-bob.  If  it  be 
of  iron  instead  of  wood,  a  link  is  used  for  a  flexible  connection. 


PUMPING.  353 

The  horizontal  beam  of  the  bob  is  about  25  feet  long,  with  a  sad- 
dle and  axle  underneath  near  the  centre.  From  the  upper  end  of 
the  king-post,  which  is  8  feet  high,  a  connecting-rod  leads  to  the 
engine.  Besides  the  braces  on  each  side  down  to  the  beam  there 
are  two  tie-rods,  taking  with  the  braces  the  tensile  and  compressive 
stresses.  All  the  members  of  the  frame  are  of  wood  or  iron,  in 
iron  shoe-castings  at  the  ends.  The  frame  stands  vertically  in  a 
pit  dug  alongside  of  the  shaft,  8  to  10  feet  deep  (Figs.  35  and 
138).  The  rocking  motion  is  communicated  to  the  bob  by  a 
connecting-rod,  or  "  pitman,"  operated  from  a  wheel  geared  to  the 
fly-wheel  shaft  of  the  engine,  the  work  of  which,  during  the  up- 
and  down-strokes,  is  somewhat  equalized  by  the  bob  and  its 
counterpoise.  The  crank-pin  can  be  set  to  different  radii  to 
alter  the  pump-stroke,  increasing  the  leverage  of  the  engine 
for  the  greater  depth  of  pumping.  The  third  apex  is  occupied 
by  a  box  full  of  iron  and  boulders,  to  counterbalance  the  ex- 
cessive weight  of  the  rod;  for  it  will  be  found  that  the  weight 
of  the  long  column-rod  of  a  strength  requisite  to  force  the 
water  up  is  much  greater  than  that  of  the  water  pumped.  A 
certain  pump,  raising  440  gallons  1690  feet  by  six  lifts  in  a 
22-inch  pipe,  has  a  balancing  weight  of  33  tons  on  the  bob. 

The  Cornish  Engine. — When  once  placed  and  its  speed  regu- 
lated, the  Cornish  pump  gives  little  trouble.  It  is  the  most 
reliable  and  also  the  most  expensive  pump  in  use.  It  has  numer- 
ous advocates  as  against  the  steam-pumps;  but  in  transplanting 
the  system  to  America  its  main  redeeming  feature — the  cataract- 
engine — was  discarded,  while  persistently  clinging  to  the  worst — 
the  cumbrous  bob  and  rod.  When  the  vertical  direct-acting 
engine  was  introduced  it  was  thought  to  be  a  great  improvement, 
because  of  the  suppression  of  the  heavy  bob;  but  it  was  soon 
discovered  to  be  a  mistake,  and  the  beam  was  quickly  reestab- 
lished, with  a  Corliss  horizontal  engine  as  the  motor.  An  engine, 
boiler,  and  fittings  complete,  with  three  15 -inch  plungers  and 
one  1 6-inch  lift-pipe,  etc.,  etc.,  for  a  600- foot  shaft,  weighed 
650,000  Ibs.,  had  a  capacity  of  800  gallons,  and  cost  in  place 
$54,000. 


MANUAL  OF   MINING. 


FlG.  138.— The  Cornish  Pump. 


All  the  foundations  about  the 
shaft  should  be  carefully  laid,  for 
the  condensed  steam  and  pump- 
water  soon  makes  the  ground 
yielding.  In  stable  ground 
heavy  beams  buried  in  the  stone 
will  suffice ;  in  ground  at  all  soft 
secure  foundation  can  only  be  ob- 
tained by  concreting  a  consider- 
able area  for  6  feet  deep,  and 
erecting  a  rigid  timber  or  brick- 
work base  several  feet  high  (Fig. 

138)- 

The  Speed  of  the  Cornish 
Pump.  —  A  variable  regulator 
enables  the  engine  somewhat  to 
change  its  speed  commensurate 
with  the  water  to  be  discharged. 
The  rod  speed  is  about  60  feet 
per  minute,  giving  six  to  ten 
strokes  of  10  or  6  feet  each.  To 
increase  the  speed  it  is  preferable 
to  increase  the  length  of  stroke 
rather  than  the  number.  The 
maximum  speed  cannot  exceed 
that  of  the  inflow,  or  the  plunger 
would  acquire  on  its  down- 
stroke  an  acceleration  before 
the  water  had  fully  entered  the 


PUMPING.  355: 

cylinder.  A  shock  ensues  which  is  detrimental  to  the  rod  and 
valves.  If  continued,  a  vibration  or  churning  occurs  with  the 
continual  pound  of  descending  plunger  upon  the  rising  water 
column.  The  speed  of  the  pump  should  be  reduced  or  the  engine 
stopped  entirely  for  a  time.  If  after  starting  it  is  again  set  up 
the  fault  may  be  with  the  valves  giving  an  obstrucled  flow. 

The  valves  should  afford  unobstructed  passage  to  the  water 
in  one  direction,  and  close  perfectly  in  the  other — Iwo  antagonis- 
tic conditions  which  can  be  attained  only  partially.  The  strain 
on  the  valves  is  enormous,  and  if  tried  too  hard  they  become 
weak,  do  not  work  properly  at  either  stroke,  and  lose  water  by 
their  "slip."  If  the  valves  are  found  to  be  free  and  the  vibra- 
tion still  continues,  the  joints  of  the  rod  are  loose  or  the  bob 
requires  resetting.  Increasing  the  mass  of  the  rod  is  sometimes, 
though  not  always,  a  remedy.  This  vibration  must  be  par- 
ticularly guarded  against  if  the  rod  is  also  to  be  utilized  for  a 
man-engine.  For  convenience  in  repairing,  the  pumping  com- 
partment should  be  provided  with  a  ladder-way,  with  plats  and 
chambers. 

Cushier's  Double-acting  Drive-rod  System. — The  use  of  a 
double-acting  pump,  retaining  therewith  the  advantages  of  the 
Cornish,  would  save  space  in  the  shaft,  the  pipes  for  a  continuous 
discharge  occupying  less  than  one  fourth  the  area  of  a  single 
discharge-pipe  and  its  rod. 

Cushier's  system  of  pumps  for  deep  mines  consists  in  having 
sets  of  two  pumps,  each  working  in  concert,  one  above  the  other, 
the  suction-  and  discharge-pipes  being  common  to  both  pumps. 

The  pumps  are  placed  at  intervals  of  about  200  feet  in  the 
shaft,  the  power  being  transmitted  directly  through  the  centre 
of  the  plungers.  The  connection  with  each  other  and  to  the 
motive  power  is  effected  by  means  of  a  steel-wire  cable  encased 
in  wood,  preserving  it  from  external  wear  as  well  as  from  rust. 
This  cable  is  fastened  to,  and  its  length  regulated  by,  shackle- 
bolts. 

The  plunger  of  the  lower  pump  in  a  set  is  double  in  area 
that  of  the  upper  one,  so  that  in  working  on  the  upper  stroke 


356  MANUAL  OF  MINING. 

one  half  the  water  raised  fills  the  chamber  of  the  upper  pump, 
the  other  half  being  forced  out  through  the  discharge-pipe  on 
the  down- stroke ;  the  upper  pump-plunger  forces  out,  in  its 
turn,  the  water  in  the  chamber,  thereby  causing  a  continuous 
delivery. 

This  form  of  pump  can  be  worked  at  any  angle,  to  any  depth, 
and  is  almost  perfectly  balanced.  The  last-named  advantage 
enables  it  to  be  connected  with,  and  worked  by,  a  direct-acting 
steam-cylinder,  and  thus  does  away  with  the  complicated  gear 
and  bob  of  the  Cornish. 

Sinking-pumps. — During  operations  of  shaft-sinking  a  double- 
acting  steam-pump  may  be  suspended  vertically  from  some  support 
by  a  chain  to  a  bale,  the  suction-inlet  being  at  the  bottom  pf 
the  pump.  By  this  suspension  it  can  be  accommodated  to  a 
varying  water-level.  A  centre-packed  plunger  is  directly  con- 
nected with  the  steam-piston,  from  which  it  receives  its  power, 
as  in  the  horizontal  pump  (see  Figs.  140  and  144).  The  sinking- 
pumps  of  the  Cameron,  Knowles,  and  Worthington  companies  are 
of  similar  pattern.  The  valves  are  absolutely  positive,  and  are 
protected  by  a  cast-iron  shield  serving  as  a  yoke  between  the 
steam  and  water  ends,  while  those  in  the  steam  end  are  cush- 
ioned to  regulate  the  strokes.  Hand-hole  plates,  with  hinged 
bolts,  allow  of  easy  repair  of  the  valves  and  the  shoes  and  dogs 
(at  the  left  of  the  pump,  Fig.  139)  of  easy  and  simple  means 
of  support. 

Sometimes  the  sinking-pumps  are  attached  to  a  sinking- 
frame  guided  in  the  shaft.  The  frame  and  pump  are  raised  as 
occasion  requires  by  chain-blocks,  winches,  or  special  hoist  at 
the  surface. 

The  Reciprocating  Pumps. — A  plunger  driven  by  steam,  air, 
or  water  pressure  in  a  cylinder  directly  connected  with  the  water- 
piston  or  plunger,  the  two  being  rigidly  coupled  and  having  a 
common  stroke  (Fig.  141),  constitutes  a  reciprocating  pump. 
It  is  always  double-acting  and  may  be  single  or  duplex,  and  is 
often  compounded  at  the  motor  end,  with,  in  some  cases,  a  con- 
densing connection. 


PUMPING. 


i3Q  _A  Sinking  pump.  FlG.  140.— Section  of  a  Sinking  pump. 


358 


MANUAL  OF  MINIX 


PUMPING. 


359 


The  column-pipes  are  smaller  in  diameter  than  for  the  Cor- 
nish system  and  the  piston  speed  is  greater,  but,  being  of  the  non- 
rotary  type,  no  advantage  can  be  taken  of  the  expansion  to  any 
great  degree.  The  water  end  may  have  a  piston  (Fig.  142)  or  a 
plunger  (Fig.  141).  The  latter  is  used  for  the  higher  heads  of 
lift. 


FlG.  142. — Packing  a  Follower  Piston. 

The  steam-pump  dispenses  with  the  cumbrous  gear,  bob,  and 
rod,  having  instead  a  small,  well-lagged  steam-pipe,  conveying 
the  power  down  from  a  surface  boiler.  Its  construction  is  sim- 
ilar to  that  of  the  air-compressor  (Fig.  126),  consisting  of  a  steam- 
cylinder  in  which  a  piston  oscillates,  moving  its  own  steam- 
valves  by  rockers  without  the  aid  of  any  rotary  appliances;  at 
.the  same  time  it  reciprocates  a  solid  plunger  centrally  in  a  water- 


MANUAL  OF  MINING. 


FlG.  144. — A  Steam  Sinking-pump  and  One  of  the  Relay 
Pumps  of  a  Forcing  System. 


FIG.  143. — A  Suction-pipe  Fitted  for  a  Condenser, 


PUMPING.  361 

cylinder,  at  each  end  of  which  is  a  set  of  double-beat  valves  of 
appropriate  construction,  open  only  so  long  as  the  water  is  being 
forced,  and  closed  with  the  aid  of  springs.  One  valve  is  removed 
in  the  figure. 

Pump-valves. — The  inlet-valves  are  opened  back  of  the 
water-piston  by  the  pressure  of  the  atmosphere,  which  forces 
water  into  the  space  thus  provided.  The  water  is  expelled 
through  valves  into  the  dischajge-pipe  on  the  return-stroke. 
At  the  other  end  of  the  water-cylinder  similar  conditions  prevail, 
the  delivery- valves  being  raised  against  the  action  of  the  springs 
which  hold  them  to  their  seat  (Fig.  141),  and  the  exhaust-valves 
being  raised  by  atmospheric  action. 

The  steam-valves  in  duplex  pumps  are  like  those  in  the  steam 
end  of  engines,  but  have  no  lap.  They  are  provided,  however, 
with  considerable  allowance  for  lost  motion,  which  prevents  one 
port  from  opening  as  soon  as  the  other  port  has  been  closed, 
at  the  end  of  the  stroke  of  the  piston.  The  valve-stem  does  not 
engage  the  valves  rigidly,  and  hence  imparts  no  motion  to  the 
valve  until  the  piston  operating  it  has  nearly  completed  its 
stroke.  The  amount  of  lost  motion  is  about  one  third  the  width 
of  the  steam-port.  If  the  piston  does  not  make  the  full  stroke, 
the  valve-chest  cover  must  be  removed  and  the  lost  motion 
increased. 

Single-cylinder  Pumps  are  often  preferred  to  duplex  pumps, 
because  they  deliver  more  water  in  proportion  to  their  size  and 
the  amount  of  steam  used.  There  is  no  interference  in  the 
action  of  the  piston  by  the  tightness  of  the  stuffing-box  on  the 
other  side,  as  in  the  duplex;  hence  there  is  no  tendency  to  reduce 
its  piston  speed.  In  the  single-cylinder  pump  the  piston  never 
starts  on  its  return  until  it  has  travelled  its  full  length  of  stroke. 
The  steam  economy  is  due  to  a  smaller  radiating  surface  for  a 
given  capacity  and  the  diminished  clearance.  The  valve-gears 
of  the  various  types  are  automatic  and  give  little  trouble,  and 
are  not  complex  in  construction.  They  usually  consist  of  some 
auxiliary  valve,  which  opens  and  closes  the  exhaust  and  steam 
admission  for  the  main  valve.  The  former  is  moved  by  the 


362  MANUAL  OF  MINING. 

piston-rod,  and  the  latter  controls  the  movement  of  the  piston. 
The  piston  of  the  single-cylinder  pump  always  makes  a  full  stroke. 

The  Plunger-pump.  —  Where  great  pressure  or  the  gritty 
nature  of  the  water  renders  the  use  of  the  single  piston  unde- 
sirable, the  water- cylinder  is  divided  in  the  centre,  and  a  pair  of 
plungers,  discharging  alternately,  work  in  the  opposite  ends, 
and  are  connected  with  yokes  and  heavy  outside  rods  to  the  steam 
piston-rod  (Fig.  145).  This  arrangement  of  external  stuffing- 
boxes  permits  of  instant  detection  of  leaks.  Strictly  speaking, 
the  combination  is  a  pumping-engine ;  but  this  term  is  custom- 
arily applied  only  to  the  double-  and  triple-expansion  engines 
used  for  city  supply. 

The  pump-cylinders  are  best  made  with  brass  linings.  The 
piston  is  packed  with  hydraulic  packing  (Fig.  142),  hemp  soaked 
in  Albany  compound  for  cold  water,  or  square-braided  cotton 
mixed  with  plumbage  for  hot  water.  Such  packing  is  held  in 
place  by  the  follower.  Rings  of  square  rubber  or  double  cup- 
leathers  are  also  sometimes  used. 

The  plunger  is  usually  hollow  and  of  such  thickness  that  it 
will  be  of  the  same  weight  as  an  equal  volume  of  water. 

An  air-chamber  is  used  on  pumps  to  maintain  an  equal  dis- 
charge. It  is  placed  at  the  highest  point  on  the  discharge  and 
permits  the  air  to  rise  to  form  a  cushion  which  maintains  uni- 
form- pressure  upon  the  column  of  rising  water.  Its  volume  is 
about  twice  the  volume  of  the  water- cylinder  of  duplex  pumps. 

A  vacuum-chamber  on  the  pump  will  keep  the  water  in  full 
motion,  and  stop  it  gradually  and  easily.  It  may  be  placed  at 
the  side  or  the  end  of  the  pump.  Its  action  is  practically  the 
reverse  of  that  of  the  air-chamber  and  facilitates  the  changing  of 
continuous  into  intermittent  motion. 

The  suction-hose  is  connected  under  the  inlet-chamber,  and 
the  discharge-pipe  to  the  surface  on  top  or  at  one  side  of  the 
outlet-chamber.  The  water-passages  are  short  and  very  direct; 
the  valves  should  be  large,  move  quickly,  and  close  tightly,  that 
little  loss  be  experienced;  otherwise  the  effective  and  suction 
powers  are  both  reduced. 


PUMPING.  363 

The  principal  difficulty  is  with  the  disposal  of  the  exhaust. 
If  turned  off  into  the  sump,  as  in  Fig.  143,  the  temperature  of 
the  mine  is  raised,  ventilation  is  injured,  and  the  timbers  ruined; 
if  carried  to  the  surface,  the  condensation  in  the  pipe  gives  trouble. 
The  best  remedy  is  to  use  a  condenser,  which  reduces  the  back 
pressure  and  increases  the  efficiency.  Jacketing  the  cylinders 
materially  contributes  to  economy  of  steam  consumption. 

The  condensation  of  the  steam  can  be  carried  into  the  suc- 
tion-pipe, as  is  the  universal  practice,  being  arranged  to  enter 
the  pipe  in  a  direction  parallel  to  the  flow  of  the  water.  This 
enables  it  to  act  like  an  injector,  and  aid  in  accelerating  the  lift 
of  the  water  (Fig.  143). 

The  Comparative  Merits  of  the  Steam-pump. — The  direct- 
acting  steam-pump  does  not  equal  the  Cornish  cataract  pumping- 
engine  in  fuel  economy.  A  fly-wheel  is  necessary  in  order  to 
secure  the  full  benefits  of  a  high  degree  of  expansion,  which, 
as  has  been  seen  in  Chapter  V,  is  not  feasible  in  one  cylinder,  be- 
cause the  resistance  (the  weight  of  water  forced  up)  is  constant 
throughout  every  stroke.  A  fly-wheel  would  distribute  the 
stearn-power  excess  of  the  first  part  of  the  stroke  to  the  latter. 
By  the  use  of  the  compound  cylinders  the  saving  of  steam-power 
may  be  fully  10  per  cent.  The  ratio  of  expansion  is  rarely  over 
3.  The  Deane  compound  pump  with  externally  packed  pistons 
is  illustrated  in  Fig.  145. 

Direct-acting  pumps  are  generally  made  as  light  as  possible, 
but  it  is  not  practicable  to  employ  the  steam  expansively.  The 
steam  is  admitted  during  the  full  stroke  of  the  plunger,  as 
is  revealed  in  the  indicator  card,  Chapter  V,  showing  the  con- 
stant pressure  required  for  the  uniform  water  resistance.  The 
pressure  may  be  changed  within  limits  of  the  boiler  supply. 

For  general  purposes  these  direct-acting  force-pumps  are  in 
universal  use.  Their  chief  feature  is  their  comparatively  high 
efficiency  at  any  speed,  slow  or  fast;  they  are  capable  of  quick 
adjustment  in  speed  and  discharge,  as  emergency  demands; 
but  they  require  close  watching,  especially  where  the  water 
is  "quick,"  or  they  may  be  drowned.  The  Cornish  pump  does 


364  MANUAL  OF  MINING. 

not  admit  of  variations  in  its  rate:  during  summer  and  winter 
it  is  run  only  a  few  hours  through  the  shift  to  empty  the  sump 
which  has  been  filling  overnight.  The  small  cost,  great  sim- 
plicity, and  ease  of  repairs  give  the  steam-pump  an  important 
advantage.  A  plant  with  boiler,  pipe,  and  fittings,  complete, 
can  be  installed  for  less  than  one  fifth  that  of  a  Cornish  outfit. 
One  for  850  gallons  per  minute,  400  feet,  cost  $15,000  in 
place.  These  pumps  have  timber  foundations,  in  a  large,  well- 


FIG.  145. — Deane  Compound  Pump  with  Externally  Packed  Pistons.. 

timbered  excavation  alongside  of  the  shaft  and  near  the  sump 
level,  which  must  practically  be  invariable.  They  are  useless 
during  sinking  without  a  sinking-pump  to  deliver  to  their  tanks. 
In  coal-mines,  and  where  the  machinery  can  be  established  for 
a  permanent  bed,  these  pumps  have  no  rival  (especially  if  com- 
pounded); whereas  in  vein-mining  the  pumping  apparatus,  and 
indeed  all  of  the  machinery,  is  under  process  of  continual  exten- 
sion. For  this  reason  in  metal-mines  the  choice  must  be  between 
a  set  of  relays  of  direct-acting  pumps  at  each  200  or  300  feet 
with  a  sinking-pump  at  the  bottom  (Fig.  144),  and  the  Cornish 
pump  with  its  several  force-stations  and  its  bottom-lift  (Fig.  138). 
The  Capacity  of  a  Pump. — In  determining  the  horse-power 
and  the  amount  of  steam  necessary  to  operate  the  pump  running 
at  full  capacity,  the  calculation  will  be  made  in  the  same 
manner  as  for  steam-engines,  bearing  in  mind  that  the  steam 


PUMPING.  365: 

cannot  be  used  expansively.  The  volume  of  steam  consumed.' 
per  minute  can  readily  be  determined  by  the  continued  product 
of  the  area  of  the  steam-piston,  the  number  of  strokes  made- 
by  all  of  the  plungers  in  a  minute,  and  the  length  of  each  stroke. 
This  product,  reduced  to  pounds  by  reference  to  the  steam-tables- 
and  divided  by  34.5,  will  determine  the  boiler  horse-power  of 
its  boiler.  Pumps  usually  have  boilers  independent  of  the  other 
steam-consumers. 

Steam-pumps  are  very  wasteful  of  steam,  the  simple  direct- 
acting  type  using  at  least  100  Ibs.  per  horse-power  per  hour;  the- 
duplex  may  consume  as  much  as  300  Ibs.  of  steam;  the  com- 
pound non-condensing  may  not  use  more  than  70  Ibs.,  and  the- 
compound  condensing  but  40  Ibs.  per  horse-power  hour. 

Let  W  =  weight  of  steam  required  by  the  pump  per  hour,  arid' 
w  =      "      ' '    i  cu.  ft.  steam  at  the  pressure  in  the  cylin- 
der. 

Then  k,  s,  and  AT"  being  the  diameter,  stroke,  and  number  of 
strokes  per  minute,  W  =o.o^4ik2sNw. 

When  it  is  found  that  a  boiler  pressure  exceeding  100  Ibs^ 
per  square  inch  is  necessary  for  the  given  lift  of  water,  either 
the  pump  or  the  boiler  pressure  must  be  increased,  or  the  height 
of  the  lift  reduced.  With  a  given  steam-pump  and  boiler  pres- 
sure it  would  be  necessary  to  employ  several  pumps  in  relays, 
each  pump  delivering  to  the  tank  above  and  thence  to  the  sur- 
face. In  Fig.  144  is  shown  the  disposition  and  arrangement 
of  such  a  set. 

The  Duty  of  the  Pump. — The  calculation  of  the  steam  and 
fuel  economy  is  easily  made;  the  necessary  elements  are  few 
in  number.  The  standard  of  comparison  of  the  work  of  a  pump 
is  its  duty — the  number  of  foot-pounds  of  work  actually  per- 
formed per  bushel  (80  Ibs.)  or  per  100  Ibs.  of  coal.  The  com- 
bustion of  one  pound  of  anthracite  gives  sufficient  heat,  theoretic- 
ally, to  do  12  million  foot-pounds  of  work.  The  ratio  between 
the  work  actually  done  and  the  power  at  the  steam  end  measures 
its  efficiency,  to  the  consideration  of  which,  in  and  about  mines, 
insufficient  attention  has  been  given  notwithstanding  its  pecuni- 


3^6  MANUAL  OF  MINING. 

any  importance.  The  duty  of  a  small  pump  is  from  7  to  15 
jnillion  foot-pounds  per  100  Ibs.  coal;  a  compound  pump  gives 
from  15  to  30  millions;  while  the  higher  types  of  pumping-engines 
•furnish  from  30  to  100  million  dynamic  units,  corresponding  to 
Ihe  consumption  per  hourly  horse-power  of  28  to  13,  13  to  6.6, 
And  6.6  to  2  Ibs.  coal  respectively.  The  consumption  of  coal  per 
liourly  horse-power  equals  198  divided  by  the  duty  (in  millions). 
A  recent  report  of  a  Wo:thington  engine  having  a  capacity  of  over 
-looo  gallons  per  minute,  against  an  equivalent  of  2000  feet  head 
of  water,  showed  a  duty  of  184  ft. -Ibs.  per  thermal  unit,  or  158,- 
«DOO,OOO  dynamic  units  per  100  Ibs.  of  coal. 

To  illustrate  the  influence  of  compounding  and  jacketing  the 
steam-cylinder,  and  of  condensing  the  exhaust,  upon  the  coal 
bills,  two  examples  will  suffice.  As  has  been  stated,  many  of 
the  collieries  pump  4000  gallons  of  water  per  ton  of  coal  hoisted. 
To  raise  this  only  300  feet  requires  the  consumption,  theoreti- 
cally, of  336  Ibs.  anthracite  for  a  daily  output  of  400  tons.  If 
the  duty  be  90  million  foot-pounds,  as  in  Cornwall,  or  20  million, 
as  with  our  average  duplex  compound  pumps,  the  aggregate 
yearly  consumption  is  675  and  3005  tons  respectively  of  anthra- 
cite, or  of  900  and  4000  tons  of  lignite. 

But  duty  is  not  the  sole  feature  of  a  piece  of  machinery: 
the  repairs  and  lubricant  accounts  and  the  durability  of  the 
plant  are  not  to  be  overlooked;  for  the  indicator  card  is  a  less 
valuable  guide  than  are  the  coal,  oil,  and  packing  bills.  More- 
over, the  inconveniences  arising  from  the  use  of  steam  under- 
ground and  those  of  the  occupation  of  a  shaft  compartment 
fcy  rods,  catches,  etc.,  must  receive  attention.  The  cost  of  pump- 
ing per  million  foot-pounds  is  not  far  from  1.6  cents  with  the 
direct-acting  pumps,  and  2.5  to  2.9  with  high-pressure  rotative 
engines. 

Pressure-regulators  for  Pumps. — A  steam-pump  should  be 
provided  with  a  speed-governor  and  a  pressure-regulator.  The 
design  is  to  place  on  the  pump  a  diaphragm  or  a  separate  piston 
•connected  by  a  spindle  with  the  valve  supplying  steam  to  the 
pump.  When  the  water  pressure  becomes  excessive  it  acts  upon 


PUMPING. 

a  balanced  diaphragm  which  opens  the  steam-pipe,  and  admits 
a  greater  supply  to  the  pump. 

A  check-valve  should  be  placed  in  the  pipe  between  the  pump 
and  the  delivery  stand-pipe,  to  prevent  the  return  of  the  water 
in  the  event  of  the  pump  stopping  operations.  A  relief-valve  or* 
the  water-pipes  will  also  prevent  any  injury  from  an  excessive- 
pressure  or  sudden  shock. 

The  Speed  of  the  Steam-pump. — Pumps  are  usually  started 
by  a  process  called  priming,  which  consists  in  removing  the  air 
from  the  barrel,  filling  the  pump  with  water  from  a  discharge- 
pipe  until  all  the  suction  space  is  charged. 

The  speed  may  be  varied,  but  is  not  advisable  beyond  100 
feet  per  minute  as  the  standard  rate.  The  water-pipe  is  calcu- 
lated for  a  flow  of  200  feet  per  minute.  The  piston  speed  is 
limited  by  that  of  the  possible  velocity  of  entry  of  the  suction* 
A  speed  in  excess  of  this  not  only  results  in  water-ram,  but  reduces 
the  capacity  of  the  pump  and  increases  the  difficulty  of  a  prompt 
seating  of  the  valves.  Any  delay  in  the  closing  of  the  valves 
results  in  a  slip  of  the  water  into  the  cylinder.  The  slip  should 
not  exceed  3  per  cent  of  the  piston  displacement.  A  large  num- 
ber of  small  valves  is  preferable  to  a  few  large  ones,  so  far  as 
the  amount  of  slip  is  concerned.  The  area  of  the  suction-valves 
should  be  40  per  cent  of  the  water-piston  area.  Their  diameter 
should  not  be  over  4  inches. 

The  Suction  Height. — The  size  of  the  suction-pipe  should 
be  as  large  and  as  short  as  possible,  that  the  pump  may  be  able 
at  high  speeds  to  obtain  as  much  water  as  it  can  deliver  with  a 
given  speed  of  plunger.  Its  size  will  depend  upon  the  length  of 
the  pipe  between  the  pump  and  the  water  level,  as  well  as  the 
difference  in  elevation  between  the  two.  The  former  determines 
the  frictional  resistance,  and  the  latter  the  maximum  velocity  of 
inlet  entering  the  pump,  due  to  the  atmospheric  pressure. 

As  the  reciprocating  piston  cannot  produce  a  vacuum  lower 
than  3  or  4  Ibs.  absolute  pressure,  the  atmosphere  cannot  lift 
the  water  higher  than  28  feet  in  the  suction  length.  To  obtain 
a  velocity  of  eatry  the  pump  must  be  placed  nearer  to  the  water 


368 


MANUAL  OF  MINING, 


PUMPING.  369 

level  than  this  suction  height.  As  it  is  desirable  to  have  a  high 
speed  of  transmission,  the  pump  should  be  lowered  to  such  a 
position  as  to  give  a  velocity  of  at  least  25  feet  per  second.  Hence 
the  maximum  suction-lift  of  14  feet  should  not  be  exceeded  if 
high  speed  is  desired. 

The  height  to  which  the  water  is  raised  is  equal  to  the  level 
of  the  suction- lift  below  the  pump  plus  the  height  to  which  it  is 
forced  above  the  pump.  This  height  multiplied  by  0.433  *s 
equal  to  the  head  acting  on  each  square  inch  of  the  water- 
piston.  The  area  of  the  water-piston  multiplied  by  this  weight 
determines  the  total  resistance.  The  steam  pressure  required 
in  the  pump  per  square  inch  of  steam-piston  is  equal  to  the 
resistance  just  determined  divided  by  the  area  of  the  steam- 
piston.  This  quantity  multiplied  by  m,  the  efficiency  of  the 
pump,  determines  the  actual  steam  pressure  necessary  to  over- 
come the  resistance  of  the  column  of  water.  Thus  a  6-inch  steam- 
piston  and  a  4-inch  water-piston  having  areas  of  28.27  and  I2-57 
sq.  in.  respectively,  with  the  water  pressure  due  to  100  feet  of 
suction  and  force  head,  or  43.3  Ibs.  per  square  inch,  the  net 
steam  pressure  theoretically  required  will  be  19.4  Ibs.  per  square 
inch;  allowing  for  a  mechanical  efficiency  of  .70,  the  steam 
pressure  then  must  be  28  Ibs.  above  that  of  exhaust. 

Motor  Fluids  other  than  Steam. — Compressed  air,  oil,  an4 
water  are  used  also  as  water  agents  in  the  driving  of  the  direct-i 
connected  pumps.  ,  The  oil-engine  is  a  simple,  self-contained, 
and  portable  motor  of  great  value.  An  oil-engine  of  high  economy 
is  illustrated  in  Fig.  147. 

Compressed  air  is  used  in  the  same  manner  as  steam  in  the. 
power  cylinders,  by  utilizing  its  elastic  property.  A  pump  driven 
by  air  cannot  be  drowned  out  as  is  a  steam-pump,  and  does 
not  require  as  much  attention  as  the  latter.  A  high-class  mod- 
ern compressed-air  plant  will  compare  very  favorably  in  com- 
mercial efficiency  with  underground  steam-pumps.  In  one 
case  280  horse-power  was  consumed  in  raising  400  gallons  per 
minute  against  a  head  of  120  Ibs.  per  square  inch,  showing  an 
efficiency  of  9  per  cent. 


37° 


MANUAL  OF  MINING. 


There  are  two  other  classes  .of  air-pumps:    the  displacement 
pump,  in  which  the  water  is  displaced  or  expelled  by  entering 


FIG.  147. — The  Diesel  Motor. 

compressed  air;    and   the  air-lift  pump,  in  which  the  water  is 
raised  by  pressure  and  the  expansive  force  of  the  air. 


PUMPING.  371 

The  Displacement  Air-pump. — Two  cylinders  are  provided 
side  by  side  with  an  air-valve  immediately  above  them.  Com- 
pressed air  is  automatically  admitted,  first  to  one  cylinder,  then 
to  the  other,  and  the  air  correspondingly  released.  The  pressure 
of  the  air  on  the  surface  of  the  water  in  the  cylinder  forces  the 
latter  through  its  discharge-valve  to  a  height  depending  on  the 
air  pressure. 

These  displacement  pumps,  or  tanks,  may  also  be  arranged 
in  multiple  series  placed  at  various  elevations  between  the  bot- 
tom and  the  top  of  the  shaft.  At  each  station  the  displacement 
is  effected  in  the  manner  indicated  by  a  pressure  sufficient  to 
raise  the  water  to  the  nearest  tank  above. 

The  Air-lift  Pump. — In  this  extension  of  the  displacement 
system  air  is  delivered  from  a  necessarily  small  pipe  into  the 
stand-pipe,  where  the  water  column  is  broken  into  short  lengths 
by  small  volumes  of  entering  air,  whose  expansive  power  over- 
comes gravitation.  The  efficiency  of  the  system  is  regarded  as 
low. 

Water-pressure  Engines  are  located  at  the  mine  below  some 
source  of  supply.  The  water  flows  under  pressure  against  the 
piston  of  the  pump,  which  communicates  the  power  to  the  rods 
of  the  Cornish  system,  or  to  a  direct-connected  plunger  which 
lifts  and  forces  the  drainage  from  the  mine  (Fig.  148).  If  the 
engine  be  placed  underground,  a  still  greater  head  is  obtained 
for  power,  the  discharge  being  effected  at  a  tunnel  level  inter- 
mediate between  the  shaft  mouth  and  the  sump  level.  This  is 
the  method  at  the  Comstock  mines.  The  great  trouble  with  these 
engines  is  in  the  valve-gear,  which  is  more  complex  and  must 
be  more  nicely  balanced  for  an  incompressible  fluid  like  water 
than  is  necessary  for  an  elastic  fluid  like  steam  or  air.  The  sudden 
shutting  of  the  valves  produces  a  not  inconsiderable  concussion 
in  the  inelastic  fluid,  which  is  entering  the  cylinder  at  a  high 
velocity.  Unless  delicately  manipulated  the  valves  fail  to  operate. 

They  will  operate  at  any  reasonable  rate  within  the  limits  of 
the  valve  movement.  The  fluid,  being  inelastic,  requires  a  very 
heavy  construction  to  resist  the  static  pressure  and  the  water^ 


372 


MANUAL  OF  MINING. 


PUMPING.  373 

ram.     Ample  time  must  be  allowed  for  pause  at  the  end  of  each 
stroke.     The  hydraulic  engines  are  expensive  plants.     They  can- 
not be  used  for  shaft-sinking,  as  the  working  pressure  cannot  be 
altered  to  meet  the  increasing  resistance.     Their  field  as  prime 
motors  at  the  surface  has  been  reduced  by  the  much  cheaper  and 
.more  durable  high-pressure  impulse-wheels,  driving  centrifugals. 
Formulae  for  Pump  Calculations. 
Let    c=  diameter  of  the  water- cylinder,  inches; 
N  =  number  of  strokes  per  minute; 
k=  diameter  of  the  steam-cylinder,  inches; 
s=  stroke,  inches; 
d=  diameter  of  the  pipe,  inches; 

h= height  of  head  lost  due  to  friction  in  the  pipes,  feet; 
L=  height  of  the  lift  including  the  suction,  ?,  feet; 
G=  discharge  in  gallons  per  minute; 
Q  =  cubic  feet  per  minute; 
/=pump  efficiency  plus  the  slip,  in  per  cent. 
Then 

G=o.oo34c2sNf;  Q  =0.0x2045 4C2sNf. 

The  pressure  per  square  inch  of  water-piston  =  0.434!,. 

The  pressure  on  the  bottom  of  the  pipe  =o.^iLd2. 

The  suction  vacuum  in  inches  =30(33.8—  /). 

Total  pressure  on  the  water-piston  =o.34i£c2. 

k*p  =0.341  (L+h}c\ 

The  work  of  the  pump  =0.0002 $$GL + /. 

The  I.H.P=o.ooo253G(L+/0-w/- 

EXAMPLES. — i.  Let  it  be  required  that  100  gallons  of  water  weighing 
833  Ibs.  be  raised  200  feet  in  one  minute.  The  number  of  ft.-lbs.  of  work 
necessary  will  be  166,600  ft.-lbs.  When  the  mechanical  efficiency  is  0.6 
the  I.H.P.  =  8.43. 

2.  Required  the  horse-power  necessary  to  furnish  60  cubic  feet  of  water 
through  400  feet  of  6-inch  pipe  and  a  suction-lift  of  4  feet. 

According  to  the  formula,  Chapter  VI,  the  head  of  friction  is  equal  to 
3  Ibs.  per  square  inch;  the  total  pressure  due  to  the  column  of  water  is 
176.6  Ibs.  per  square  inch;  the  velocity  through  a  6-inch  pipe  is  equal  to 
300  feet  per  minute;  whence  the  work  done  is  179.6x28.3  X3oo  =  46.1  H. P. 
The  steam  end  must  be  capable  of  furnishing  more  power  than  this,  and  if 


374  MANUAL   OF  MINING 

the  initial  steam  pressure  be  ico  Ibs.  absolute,  with  a  back  pressure  of 
18  Ibs.,  the  diameters  of  the  steam  and  water  cylinders  with  ^=0.50  and 
a  piston  speed  of  200  feet  per  minute,  will  be  16  and  8  inches  respectively. 
The  pump  receives  its  supply  from  a  boiler  which  is  independent  of  those 
feeding  the  hoisting-engines,  because  the  intermittent  work  of  the  latter 
causes  such  changes  in  steam  pressure  as  to  seriously  affect  the  speed  of 
the  pump. 

3.  A  mine  delivers  ifoo  gallons  per  minute.     The  depth  of  the  shaft  is  468 
feet.     Required  the  size  of   the  pump-cylinders  under  a  boiler  pressure  of 
100  Ibs.  (gauge)  and  a  $  cut-off,  back  pressure  being  16  inches  of  mercury. 
Efficiency,  60  per  cent. 

As  the  ordinary  piston  speed  is  200  feet  per  minute,  the  discharge  of  4.227 
cubic  feet  per  second  may  be  delivered  at  the  same  speed  in  the  pipe,  which  is 
then  of  a  diameter  of  15  inches;  or  if  a  jo-inch  discharge-pipe  be  employed, 
the  velocity  therein  would  be  464  feet.  The  loss  of  head  would  be  respectively 
1.56  and  11.86  feet.  Total  head  being  then  470  or  480  feet,  the  work  of  rais- 
ing the  1900  gallons  would  be  7,441,500  and  7,599,840  ft.-lbs.  respectively. 
The  two  steam  ends  must  be  capable  of  12,402,516  and  12,666,400  ft.-lbs. 
From  the  table  in  Chapter  IV,  the  mean  pressure  corresponding  to  a  cut-off 
of  J  is  0.726  for  i  lb.,  and  83.27  Ibs  for  114.7  absolute. 

Let  the  average  effective  pressure  be  64  Ibs.,  then  the  diameter  of  each 
steam  end  should  be  \2\  inches.  Assuming  a  stroke  of  24  inches,  each  stroke 
represents  2.53  cubic  feet,  and  the  diameter  of  each  water-cylinder  would 
be  lof  inches.  If  the  minimum  effective  pressures  upon  the  two  steam- 
pistons  be  taken  (25  Ibs.  per  square  inch  on  one  and  64  Ibs.  on  the  other), 
the  diameter  of  the  water-cylinders  should  not  exceed  i\  inches.  This 
discussion  neglects  the  inertia  of  reciprocating  parts. 

If  the  discharge-pipe  be  assumed  at  6  inches  in  diameter,  it  would  entail  a 
loss  of  head  of  151.47  feet,  requiring  475  horse-power. 

The  ratio  between  the  diameters  of  the  two  ends  of  a  steam-pump  is  about 
i  :  2  for  the  smaller  sizes,  and  the  steam  end  is  three  times  that  of  the  water 
end  in  the  large  sizes. 

4.  What  volume  could  be  raised  by  a  double  acting  steam-pump  having 
water-cylinders  8  inches  in  diameter,  and  the  steam-cylinders  18  inches,  with  a 
2-foot  stroke  ?     Piston  speed  200  feet  per  minute.  139.4  cubic  feet. 

5.  What  should  be  the  effective  steam  pressure  to  discharge  the  water, 
assuming  an  efficiency  of  50  per  cent  and  a  shaft  400  feet  deep?    Assuming 
a  discharge-pipe  of  5  inches  diameter,  p  is  83.7  Ibs.  per  square  inch,  the  loss 
of  head  being  88.22  feet. 

Power-driven  Pumps. — Where  it  is  not  possible  or  desirable 
to  connect  the  water-plungers  directly  with  an  engine,  power 
may  be  communicated  to  a  pulley  on  a  rotary  shaft  by  a  belt 


PUMPING.  375 

or  wire  rope,  or  to  an  electric  motor.  If  the  distance  between 
the  engine  and  the  pump  is  very  great,  electricity  is  the  agent 
for  the  motor.  If  the  distance  is  slight,  a  belt  is  employed,  with 
or  without  gearing,  to  reduce  the  speed  of  the  motor  to  that 
of  the  pump,  which  is  always  low.  Power- pumps  are  built 
with  more  than  one  trunk-cylinder  whose  pistons  are  single- 
acting,  the  shaft  being  a  multiple  crank.  With  three  plungers 
the  cranks  are  60°  apart,  with  four  they  are  90°,  etc.  Trip- 
levers  can  be  attached  to  relief-valves,  so  that,  when  the  pres- 
sure becomes  excessive,  they  will  permit  the  water  to  escape  and 
simultaneously  close  the  suction-valve. 

The  cylinders  are  generally  vertical,  though  in  the  Riedler 
pump  they  are  horizontal.  Double-acting  duplex  or  triplex 
pumps  can  also  be  had,  of  the  power-driven  type,  which  are  used 
for  pressures  up  to  700  Ibs.  in  the  anthracite  region. 

The  power-pump  is  of  practical  utility  and  simple,  occupying 
less  floor  space  for  a  given  capacity  than  the  direct-acting.  It 
is  of  low  speed  and  is  therefore  geared  down  from  the  driving 
source.  Its  loss  by  friction  is  a  little  greater  than  in  the  direct- 
connected  pump.  Where  fuel  is  expensive  it  has  a  distinct 
advantage  in  the  point  of  steam  economy,  for  the  direct-acting 
type  cannot  use  steam  expansively.  The  power-driven  pump 
may  be  operated  by  a  modern  automatic  cut-off  engine  and 
develop  power  with  less  steam  than  the  direct-acting.  The  latter 
cannot  use  the  steam  expansively,  as  is  possible  with  the  engine 
and  a  fly-wheel-crank-driven  pump. 

The  Riedler  Pump  has  mechanically  operated  valves  which 
will  open  and  close  at  high  rotative  speed.  Its  capacity  is  large 
for  the  space  it  occupies  and  it  requires  little  foundation.  At 
its  maximum  speed  of  300  r.p.m.  it  can  raise  water  1000  feet. 
The  single-acting  plungers  are  installed  two  or  three  on  a  line,  to 
lighten  the  work  and  economize  power. 

Electrically  Driven  Pumps  are  necessarily  of  the  geared-crank 
type,  for  this  power  is  not  adapted  to  reciprocation.  Either 
continuous-  or  alternating-current  motors  are  suited  for  them. 
The  main  requirements  are  a  practically  constant  moment  or 


376 


MANUAL   OF  MINING. 


torque,  and  a  nearly  constant  speed.  With  a  continuous  current 
the  compound- wound  motor  is  employed;  if  a  variable  speed  is 
demanded,  a  rheostat  in  the  shunt-field  will  suffice.  Its  advan- 
tage over  an  ordinary  shunt-motor  is  that  its  series-coils  obviate 
the  wide  variation  of  current  which  would  occur  with  the  latter 
when  passing  through  the  different  points  of  the  pump  cycle. 
It  will  not  race  if  the  pump  happens  to  lose  its  water.  Prefer- 
ably, it  should  not  be  enclosed.  In  the  shunt-wound  motor 
the  field  resistance  will  rise  as  the  enclosed  machine  gets  hot, 
thus  causing  higher  speed  and  armature  current.  The  diffi- 
culty can  be  avoided  by  a  few  additional  turns  of  series  winding. 


FIG.  149. — Electric  Power-pump. 

With  the  alternating  current  the  short-circuited  squirrel-cage 
type  of  armature  induction-motor,  designed  for  any  desired  turn- 
ing moment  and  operated  at  a  constant  speed,  is  well  adapted 
to  drive  the  pumps. 

The  leakage  loss  of  electric  pumps  is  small  and  their  friction 
loss  low,  while  the  expense  of  maintenance  and  repairs  is  a 


PUMPING.  377 

minimum.  One  compartment  may  be  saved  in  the  shaft,  since 
the  wires  and  the  discharge-pipes  can  be  placed  in  either  of  the 
hoistways  without  interference.  In  common  with  the  other 
motor-driven  pumps,  it  is  portable.  It  is  started  and  stopped 
easily  by  the  ordinary  switch  with  an  automatic  release-switch. 

Like  the  power  end  of  the  air-  and  water-driven  pumps,  the 
motor  may  be  operated  even  under  water,  for  it  can  be  com- 
pletely housed  with  its  sinking-pump. 

Electrically  driven  pumps  require  automatic  switches  which, 
when  the  water  level  reaches  a  given  point,  engage  a  lift  that 
moves  the  switch  and  opens  or  closes  the  circuit,  as  may  be 
desired. 

Rotary  Pumps  are  employed  for  lifting  and  forcing  water 
against  low  head.  This  type  of  pump  is  compact  and  self-con- 
tained and  will  deliver  more  water  for  a  given  weight  and  space 
occupied  than  will  reciprocating  pumps.  It  is  usually  driven 
by  belting  or  wheel  gearing.  Rotary  pumps  may  be  divided 
into  several  classes,  according  to  the  forms  of  the  pistons  or  im- 
pellers, or  according  to  the  arrangement  of  the  butments.  Two 
rotary  impellers  receive,  in  the  space  between  them  from  the 
suction  below,  a  volume  of  water  which  is  carried  with  the  evolu- 
tion of  the  impellers  around  to  the  upper  side  of  the  pump. 
The  butments  receive  the  force  of  the  water,  and  prevent  the 
latter  from  being  carried  around  the  cylinder,  thus  compelling 
it  to  enter  the  delivery,  where  it  is  discharged.  The  Sturtevant 
blower  has  two  rotary  pieces,  and  the  Root  three  lobes. 

In  some  pumps  the  butments  are  movable  and  are  arranged 
to  be  pushed  back,  as  they  revolve,  to  allow  the  piston  to  pass  a 
given  point;  in  others  the  pistons  give  way  when  passing  fixed 
butments;  and  in  others  the  pistons  are  fitted  with  a  movable 
wing,  as  in  the  fan,  which  slides  radially  in  and  out  when  pass- 
ing the  butments.  These  pumps  have  no  packing  or  springs, 
are  quite  durable,  but  have  a  tendency  to  become  noisy  as  the 
gearing  wears. 

The  Centrifugal  Pump. — This  is  the  cheapest  and,  under 
certain  conditions,  the  most  efficient  of  the  rotary  variety  of 


378  MANUAL  OF  MINING. 

pumps  where  the  height  of  lift  is  moderate  and  large  volumes 
are  to  be  moved.  Like  all  of  the  rotary  type,  it  handles  sandy 
or  muddy  waters  with  facility,  and  has  a  very  short  suction  and 
only  a  moderate  height  to  which  it  can  raise  water.  It  is  a  fan 
with  a  number  of  blades  attached  to  a  shaft,  and  is  turned  by  an 
electric  motor  or  steam-turbine  at  500  to  800  r.p.m.  The  inlet 


FIG.  150. — Section  of  a  Centrifugal  Pump. 

may  be  on  one  side  only,  or  on  both  (Fig.  151).  With  the  latter 
the  inlets  and  blades  may  be  kept  down  to  smaller  sizes  for  an 
equal  capacity.  The  water  enters  at  the  centre  and  is  delivered 
at  the  circumference  by  the  high  speed  of  the  blades.  The 
blades  are  convex,  radial,  or  concave.  The  former  require  a 
higher  rate  of  velocity  for  a  given  lift  than  do  radial  blades. 
Their  inner  ends  are  curved  forward  slightly  to  scoop  up  the 
water  with  a  minimum  of  shock.  The  inlet  velocity  should  not 
exceed  3  feet  per  second.  The  pump  is  primed  and  started 


PUMPING. 


379 


by  some  form  of  steam-  injector  which  expels  the    air  and  fills 
it  with  water. 

The  Theory  of  the  Centrifugal  Pump.  —  The  water  is  driven 
through  the  fan  partly  by  the  pressure  of  its  blades  and  partly 
by  centrifugal  force,  and  the  water  will  escape  with  that  force 
at  c  (Fig.  150),  but  with  little  rotary  motion,  and  will  rise  to  a 
height,  h,  corresponding  to  the  centrifugal  force.  If  there  be 
no  unavoidable  efficiency  losses,  the  relation  between  the  total 
lift,  h,  and  the  peripheral  velocity,  V,  is  expressed  by 


If  the  outlet  for  the  water  be  of  proper  size  and  flared  to 
reduce  the  velocity  of  discharge  gradually,  the  energy  imparted 
to  it  by  the  blades  is  converted  into  pressure  head.  With  that 
due  to  centrifugal  force,  the  head  ideally  obtainable  from  a  periph- 


FlG.  151. — Double-inlet  Centrifugal  Pump. 

eral  velocity,  F,  is  2h  (Fig.  150).     The  usually  imperfect  pump, 
however,  shows  a  relation  between  velocity  and  head  of 


in  which  u  is  the  velocity  in  the  discharge-pipe.     The  head,  h,  is 
only  about  60  per  cent  of  that  theoretically  due. 


380  MANUAL  OF  MINING. 

To  increase  the  discharge  for  a  given  head  means  loss  of 
efficiency,  because  the  speed  of  the  pump  and  that  of  the  water 
issuing  from  the  pipe  are  increased  unnecessarily;  a  reduction 
of  velocity  is  also  attended  with  loss  by  fluid  friction.  Where 
efficiency  is  desired  the  design  must  be  just  suited  to  the  precise 
conditions  under  which  it  is  to  operate.  The  first  cost  is  low 
and  its  efficiency  increases  to  70  per  cent  at  about  40  feet  of 
lift,  beyond  which  height  the  efficiency  again  diminishes,  until 
at  a  height  of  100  feet  the  efficiency  is  but  40  per  cent. 
Let  A/"=r.p.m.  of  the  wheel; 

H  =  height  of  delivery  in  feet; 
D  =  diameter  of  wheel  in  feet; 

K=  constant;    153  for  small  pumps  and   187  for  large 
pumps; 


then  N 

EXAMPLE.  —  A  centrifugal  pump  discharges  180  cubic  feet  per  minute  at  a 
peripheral  velocity  of  30  feet  per  second.  Required  the  horse-power  to 
drive  the  pump.  The  peripheral  velocity  of  the  water  is  25  feet  per  second. 

Then  momentum  lost  by  the  wheel  per  second=-  —  —X  25  =  145.  5  Ibs. 

The  work  done=  145.5X30=4365  ft.-lbs.  per  second=8  horse-power. 
The  efficiency  being  assumed  at  70  per  cent,  the  horse-power  applied  to 
the  pump  is  11.5. 

4365  +  3X62.5  =  23.3  ft.-lbs.  energy. 

This  would  raise  a  pound  of  water  23.3  feet. 

The  Compound  Centrifugal  Pump.  —  These  pumps  are  essen- 
tially of  a  high-speed  type  with  low  head  of  lift.  A  high  lift  is 
attained  only  by  increasing  the  rotary  speed.  If,  however,  two 
or  three  wheels,  each  in  its  own  chamber,  are  mounted  in  series 
on  the  same  shaft,  the  lift  of  each  single  one  is  multiplied,  while 
yet  keeping  the  speed  within  moderate  bounds  (Fig.  152). 

In  such  a  multi-stage  pump  the  water  from  the  discharged 
chamber  of  the  first  impeller  is  led  back  to  the  suction-point  of 
the  next  impeller  through  channels  in  the  pump-casing.  This  is 


PUMPING. 


381 


repeated  as  often  as  there  are  impellers.  Each  pump  of  the 
series  raises  the  pressure  head  from  that  of  its  suction.  A  three- 
stage  pump,  with  each  member  capable  of  100  feet  of  head,  will 
then  deliver  water  to  a  height  of  300  feet  with,  of  course,  a  corre- 
sponding increase  of  motor-power.  These  are  much  used  in 
placer-mining  service  instead  of  a  long  pipe-line  from  a  distant 
lofty  reservoir. 

The  Efficiency  of  the  Centrifugal  Pump. — The  efficiency  of 
these  pumps  drops  rapidly  when  the  difference  between  the  out- 
let and  the  inlet  pressure  heads  exceeds  60  feet,  and  the  efficiency 
of  a  multi-stage  pump  is  about  the  same  (0.85)  as  that  of  a  single 


FlG.  152. — A  Three-stage  Pump. 


pump  of  the  same  construction,  but  the  advantage  of  the  former 
lies  in  the  reduced  velocity  for  the  given  height  of  lift.  On  the 
same  work  and  within  reasonable  limits,  the  multi-stage  cen- 
trifugal is  slightly  more  efficient  than  the  single  pump,  due  to 
the  decrease  in  frictional  loss  attaining  the  reduced  rotary  speed. 
The  steam-turbine,  or  the  electric  motor,  is  capable  of  direct 
connection  with  the  centrifugal  pump,  and  this  combination 


MANUAL   OF  MINING. 


compares  very  favorably  in  steam  consumption  with  the  direct- 
acting  type. 


HORSE-POWERS  AND  FUEL  REQUIRED  FOR  Two-  AND  THREE-STAGE  CENTRIFU- 
GAL PUMPS  FOR  HYDRAULIC  MINING  OR  FOR  PUMPING  PURPOSES. 


Capacity. 

Horse- 

Pounds 

Miner's 
Inches. 

Gallons 
per 
Minute. 

Cubic 
Feet  per 
Second. 

at  70 
Per 
Cent. 

Diam- 
eter. 

Coal, 

I  2 

Hours. 

Short 
Tons. 

Barrels 
Oil. 

§sa 

5° 

562 

1.2 

41 

5" 

2,460 

1.2 

4-4 

2 

75 

843 

1.8 

62 

6" 

3,690 

1.8 

6.7 

3 

100 

1125 

2-5 

82 

6" 

4,920 

2.4 

8.9 

4 

125 

1406 

3-i 

i°3 

7" 

6,150 

3-i 

ii.  I 

5 

150 

1687 

3-7 

124 

8" 

7,380 

3-7 

13-3 

6 

175 

1968 

4-4 

J45 

9" 

8,610 

4-3 

15-5 

7 

200 

2250 

5-o 

164 

10" 

9,840 

4.9 

17.8 

8 

225 

2531 

5-6 

185 

10" 

11,070 

5-5 

20.  o 

9 

250 

2812 

6.2 

206 

10" 

12,300 

6.2 

22.2 

10 

275 

3°93 

6.8 

227 

12" 

13,500 

6.8 

24.4 

ii 

300 

3375 

7-4 

246 

12" 

14,760 

7-4 

26.6 

12 

35° 

3937 

8.6 

267 

12" 

17,220 

8.6 

28.9 

14 

400 

4500 

9.9 

328 

14" 

19,680 

9.8 

35-5 

16 

500 

5625 

12.4 

410 

14" 

24,600 

12.3 

44-4 

18 

A  single-stage  pump  making  1500  revolutions  per  minute, 
with  55  horse-power  steam-turbine,  delivered  1700  gallons  per 
minute  with  a  lift  of  100  feet.  The  diameter  of  the  plunger  is 
13!  inches.  A  two-stage  centrifugal  pump,  9  inches  in  diameter, 
was  making  2000  revolutions  per  minute  with  a  Dela  valve  steam- 
turbine  running  at  20,000  r.p.m.,  and  raised  250  gallons  per 
minute  700  feet.  The  tests  of  these  showed  the  efficiency  of 
the  wheel  and  turbine  to  be  75  per  cent  in  the  first  case  be- 
tween the  limits  of  1400  and  1800  gallons  per  minute.  The 
duty  of  the  pump  per  thousand  pounds  of  steam  operated 
with  a  condenser  was  nearly  62,000,000  ft.-lbs.  The  two-stage 
pump  showed  a  duty  at  250  gallons  per  minute  of  48,800,000 
ft.-lbs.  per  thousand  pounds  of  steam.  Forty-two  pounds  of 
steam  were  used  per  water  horse-power. 

A  test  of  a  single-stage  pump  direct-connected  electric  motor 
of  20  horse-power  delivering  1200  gallons  per  minute  at  2000 


PUMPING.  383 

revolutions  for  a  lift  of  45  feet  showed  an  efficiency  of  76  per  cent. 
The  diameter  of  its  wheel  was  8.3  inches. 

EXAMPLES. — i.  What  will  be  the  size  of  the  discharge-pipe,  d,  the  wheel 
and  its  rate  of  revolution,  to  deliver  1000  gallons  per  minute  for  a  height  of 
50  feet? 

d=  0.25 v^— £=0.25X31.6=  7.8  inches  pipe; 

wheel  diameter=  2  X  7.8=  15.6  inches; 
and 

JV=  187X^-^=880.67  r.p.m. 

2.  Required  the  size  of    a  three-throw  electric   pump,  the  horse-power 
and  current  at  550  volts  for  a  motor  of  90  per  cent  efficiency,  to  deliver  240 
gallons  300  feet  high.       Speed  of  piston  60  feet  per  minute.     Let  }=o.gj. 
Then  £=5.86  inches  and  the  stroke  i|c=8  inches. 

The  work  of  pumping  is  240X8^X300=590,000  ft.-lbs.  Assuming  the 
inertia,  friction,  etc.,  to  be  50  per  cent,  i  requires  36  H.P.  at  the  motor-shaft. 
With  90  per  cent  efficiency,  this  =29,840  watts,  which  at  550  volts=54-3 
amperes. 

3.  It  is  desired  to  raise  90  gallons  of  water  per  minute  up  an  incline 
of  i  in  5  which  is  1200  feet  long.     Assuming  a  discharge-velocity  of  3  feet 
per  second,  what  will  be  the  size  of  the  pipes  and  of  the  pumps  and  motor? 
If  the  three-throw  pump  is  employed,  each  pump  raises  30  gallons;   with  20 
per  cent  of  slip,  36  gallons  must  be  raised  about  60  feet  per  minute,  with  a 
wheel-diameter  of  4.25".     At  a  stroke  of  8  inches  and  90  revolutions  per 
minute,  the  diameter  o    the  discharge-pipe  will  be  3.875  inches  to  deliver 
90X0.16  cubic  feet  per  minute.   A  pipe  of  4  inches  diameter  will  lose,  accord- 
ing to  the  formula  in  Chapter  VI,  1.22  feet  of  head  for  each  100  feet  of  length 
at  the  assumed  velocity.     The  loss  of  head  due  to  friction  is  then  14.6  feet. 
The  work  of  lifting  the  water,  including  the  above  loss,  is  190,980  ft.-lbs.  per 
minute  and  the  horse-power  is  nearly  6.      Allowing  a  friction  of  gear  and 
starting  of  50  per  cent,  12  B.H.P  are  required.   The  motor  will  run  at  630  rev- 
olutions per  minute  with  a  gear  of  7  to  i .     If  the  latter  has  an  efficiency  of  80 
per  cent  and  receives  a  current  at  380  volts,  24.5  amperes  will  be  required 
to  deliver  the  10  B.H.P.  at  380  volts.     The  wire  for  this  current  must  be 
19/19  and  the  drop  in  voltage  for  the  total  length  of  wire  down  the  slope 
will  be  24.5X1.786,  which  equals  44  volts  per  mile  and  20  volts  for  2400  feet. 
The  voltage  at  the  bottom  is  therefore  400. 

The  Pulsometer. — This  useful  apparatus  raises  sandy  or 
acid  waters  to  limited  heights  where  economy  of  fuel  is  less  in? 


384  MANUAL  OF  MINING. 

portant  than  quickness  of  installation.  It  is  perfectly  free  from 
risks  of  breakdowns,  and  is  employed 
in  open  works.  A  A  (Fig.  153)  are  two 
chambers  into  which  steam  enters  accord- 
ing to  the  position  of  the  ball,  C,  which 
oscillates  from  one  side  to  another  of 
their  necks.  Through  the  inlet  passage, 
Z>,  the  water  enters  into  A  by  opening 
the  valves,  E.  H  is  a  delivery  passage 
communicating  with  each  chamber 
through  openings  and  valves,  G. 

Steam  enters  at  the  top,  passes  into  a 
chamber   uncovered    by   the    ball-valve, 

FIG.  153.— The  Pulsometer.   and  presses  upon  the  surface  of  the  water, 

forcing  it  down  and  out  through  the  discharge-valves  into  H. 
When  the  water-line  falls  below  the  discharge  outlet,  the  steam 
above  condenses,  a  partial  vacuum  is  formed,  and  its  pressure 
suddenly  falls  again.  Meanwhile,  with  the  collapse  of  the  steam, 
the  ball- valve  is  thrown  to  close  that  chamber,  and  admits  steam 
to  the  other  side.  Here  the  water  is  expelled,  in  a  similar  manner, 
out  through  H ,  while  at  the  same  time  water  has  entered  through 
D  to  be  expelled  from  the  other  side. 

The  limit  of  lift  is  about  30  feet,  and  the  capacity  of  some 
pulsometers  is  often  1000  gallons  per  minute. 

Siphons. — Though  not  a  water- raising  appliance  in  the  proper 
sense  of  the  term,  since  water  by  this  apparatus  can  be  lifted 
or  transported  over  an  eminence  not  exceeding  28  feet  in  height 
and  discharged  on  the  other  side  of  it  at  a  level  lower  than  that 
of  the  supply,  a  siphon  can  find  application  for  forming  a  water 
communication  over  a  slight  elevation  between  two  distant  points. 
This  height  will  be  further  reduced  by  an  amount  necessary  to 
cause  the  required  velocity  of  flow  and  to  overcome  the  frictional 
resistance.  Formulae  in  Chapter  VI  will  determine  the  losses  due 
to  friction  in  the  given  case,  L  and  d  being  given. 

When  the  siphon  is  not  too  long,  and  when  the  acceleration 
head  is  sufficient  to  give  the  water  a  considerable  velocity,  the 


PUMPING.  385 

air,  entrained  by  the  rapid  current,  may  be  carried  out  at  the 
end  of  the  discharge  branch,  if  the  latter  is  not  too  steep.  In 
most  cases,  however,  it  is  necessary  to  provide  artificial  means  to 
remove  the  accumulated  gases,  either  periodically  or  continu- 
ously. A  hand-pump  is  usually  employed  for  this  purpose,  its 
suction  being  connected  to  the  highest  point  of  the  siphon,  and 
operated  as  occasion  requires. 

REFERENCES. 

Compressed-air  Pumps,  111.  Min.  Inst.,  I,  100;  Pumps,  Breakdowns,  Coll. 
Guard.,  Nov.  13,  1896,  567;  Mine  Pumps,  Breakdowns,  Coll.  Guard.,  Nov. 
13,  1896,  930;  Pump  Valves  and  Their  Care,  Dugald  Baird,  Coll.  Eng.,  1893, 
221;  Mine  Drainage  and  Pumps,  Hans  B.  Behr,  California  State  Mining  Bureau 
Bull.  9,  210;  Pumps,  Description,  Pumps  in  England,  Gears,  Strokes,  etc., 
Coll.  Guard.,  May  7,  1897,  855;  Pumps,  Calculations  of,  Bjorling,  Coll.  Man- 
ager, Nov.  20,  1896,  567;  Pump  Cylinders,  Determining  Strength  of,  Kettell, 
Am.  Soc.  Mech.  Eng.,  XVII;  Careof  Pumps,  M.  &  M.,  Vol.  XXIV,  323; 
The  Engineer,  July  1902;  Sizes  and  Weights  of  Tubing,  O.  J.  Edwards, 
E.  &  M.  Jour.,  LVIII,  387;  Pumping  by  Compressed  Air,  Frank  Richards, 
E.  &  M.  Jour.,  LIX,  3  4;  Pneumatic  Pumping  Plant,  Proc.  Am.  ?oc.  C.  E., 
XXX,  445;  Hoisting  Water  Tandem  Tanks,  J.  W.  Bowden,  Amer.  Inst.  M. 
E.,  XX,  343;  Mine  Pump,  Breakdowns,  Michael  Longridge,  Coll.  Guard., 
Nov.  1896,  930;  The  Contest  with  Water  at  the  Comstock  Mines,  D.  L.  Lord, 
U.  S.  G.  S.,  IV,  230;  Tapping  Drowned  Workings,  W.  B.  Wilson,  Jr.,  Coll. 
Guard.,  Feb.  14,  1902;  Water  Hoisting  in  Anthracite  Mines,  M.  &  M.,  Vol. 
XXIII,  390. 

Centrifugal  Pumps  for  Mine  Work,  W.  R.  Crane,  M.  &  M.,  June  1902; 
Tests  of  Centrifugal  Pumps,  Engineer,  Chicago,  June  15,  1904;  Theory  of 
Centrifugal  Pump,  Am.  Soc.  C.  E.,  Jan.  1904;  and  Proc.  Am.  Soc.  C.  E., 
May  1903,  and  Vol.  XXIX,  839. 

Determining  C  O  in  Mine  Air,  Coll.  Guard.,  Vol.  LXXXIV,  1162;  Grav- 
itation of  Fire  Damp,  Coll.  Guard.,  LXXXI,  124. 


CHAPTER  XH. 

MINE-GASES. 

Ventilation  as  an  Economic  Proposition. — The  ventilation  of 
a  mine  is  a  matter  of  very  great  concern,  not  only  from  a  humani- 
tarian standpoint,  but  from  an  economic  point  of  view.  No 
shaft  or  tunnel  can  be  carried  more  than  200  feet  beyond  an 
opening  without  some  special  means  of  stirring  and  freshening 
the  stagnant  air.  The  men  should  not  be  compelled  to  work 
in  the  hot  atmosphere  of  a  stove  or  room  vitiated  by  the  variety 
of  gases  given  off  from  and  by  coal,  powder,  lamps,  respiration, 
rotting  timbers,  and  decomposing  ore.  These  gases  cannot 
support  combustion,  nor  can  they  be  inhaled  with  impunity; 
and  such  an  atmosphere  is  unfit  for  respiration,  being  deficient 
in  oxygen,  as  well  as  by  reason  of  the  presence  of  these  gases. 
In  coal-mines  additional  peril  accompanies  some  of  these  gases, 
which  with  the  air  form  explosive  mixtures,  which,  bursting 
into  flame,  destroy  everything  in  their  path  and  emit  dense  vol- 
umes of  poisonous  fumes  that  are  fatal  to  all  who  have  escaped 
the  shock. 

Recognizing  the  pecuniary  value  of  the  life  and  energy  of  a 
miner,  the  statutes  are  becoming  more  and  more  rigorous  in 
the  insistence  of  safety  and  hygienic  measures.  Methods  are 
specified  for  rendering  the  noxious  gases  harmless,  and  officials 
are  given  sufficient  authority  to  suggest  needed  improvements 
and  to  punish  non-compliance.  Not  only  are  the  miners  benefited 
by  the  diminishment  of  risk,  but  also  the  operators,  who  profit 
in  the  increased  energy  of  men  working  under  favorable  circum- 
stances. The  illumination  is  better,  smoke  clears  away  quicker, 
and  the  men  are  capable  of  a  full  day's  work. 

386 


MINE-GASES.  387 

Comparison  of  Metal-  and  Coal-mines. — There  is  a  marked 
contrast  between  the  requirements  of  vein-mines  and  those  of 
coal  or  other  mines  in  flat  beds,  which  latter  usually  have  two 
shafts  connected  by  a  labyrinth  of  workings  on  nearly  one  level. 
The  ventilation  of  metal-mines  presents  by  no  means  the  same 
difficulty  as  that  of  gaseous  collieries,  since  as  a  rule  the  former 
class  of  mines  requires  but  a  small  supply  of  fresh  air,  abundant 
enough  for  the  health,  comfort,  and  effective  work  of  the  men 
and  for  the  removal  of  the  dead  air  vitiated  by  various  causes. 
Trust  is  placed  in  natural  means  of  circulating  the  air  by  the 
winze  communication  of  the  different  levels.  This  is  certainly 
inadequate,  and  the  lower  mortality  in  well-ventilated  coal- 
mines is  doubtless  due  to  their  better  ventilation.  In  coal-mines 
large  volumes  of  fresh  air,  additional  to  that  required  for  the 
men,  are  necessary  to  carry  off  the  fire-damp,  choke-damp,  and 
other  gases,  both  noxious  and  inflammable,  though  no  serious 
results  may  occur  from  their  presence  in  large  quantities.  These 
gases  are  continually  being  evolved  from  the  coal  and  constitute 
an  ever-present  danger  to  coal  operations,  but  do  not  threaten 
metal-miners.  Nevertheless,  the  latter  should  remember  that 
the  inflammable  gases  of  coal-mines,  from  which  they  are  com- 
paratively free,  are  not  the  only  ones  to  be  guarded  against.  The 
cloud  of  dust  formed  by  blasting  the  mineral,  and  the  carbonic 
acid  gas,  are  dangers  equally  insidious  to  the  health  of  those 
inhaling  it,  which  cause  as  many  deaths  as  do  the  small  explo- 
sions which  occur  in  coal-mines. 

Gases  in  Coal-mines. — In  addition  to  the  gases  found  in  a 
normal  atmosphere,  there  exist  in  mines  several  gases  which 
are  always  the  result  of  decomposition  and  combustion.  Pur- 
suing the  nomenclature  of  early  mining,  they  are  designated 
as  "  damps."  For  example,  carbonic  acid,  CO2,  is  known  as 
choke-damp;  carbonic  oxide,  CO,  as  white  damp;  sulphuretted 
hydrogen,  H2S,  as  stink-damp;  carburetted  hydrogen  or  marsh- 
gas,  CH4,  as  fire-damp  or  fulminating  damp;  air  vitiated  by 
breathing,  having  therefore  a  deficiency  of  oxygen,  as  black 
damp;  and  the  residual  gases  of  an  explosion,  after-damp.  In 


388  MANUAL  OF  MINIXG. 

metal-mines  carbonic  acid  and  sulphuretted  hydrogen  are  the 
only  gases  met  with.  Both  are  heavier  than  air  and  naturally 
will  be  found  near  the  floor  of  the  workings.  Coal-mines,  how- 
ever, are  troubled  with  additions  and  emanations,  sudden  or  con- 
tinual, of  carbonic  oxide  and  marsh-gas,  which  two,  being  lighter 
than  air,  are  to  be  looked  for  in  the  upper  portions  of  the  work- 
ings, near  the  roof. 

Composition  of  the  Atmosphere. — Every  hundred  grains  of 
the  atmosphere  contain  76.84  grains  of  nitrogen,  23.10  grains  of 
oxygen,  and  0.06  grain  of  carbonic  acid,  occupying  respectively 
79.02  per  cent,  20.94  per  cent,  and  0.04  per  cent  of  the  original 
volume.  These  gases  are  not  in  chemical  combination,  but  as  a 
mechanical  mixture  from  which  the  oxygen  may  be  extracted 
under  given  conditions. 

Oxygen,  which  is  colorless,  odorless,  and  transparent,  unites 
with  all  other  chemical  elements,  forming  various  chemical  com- 
pounds. All  the  ordinary  phenomena  of  light,  heat,  and  fire 
are  the  result  of  the  union  of  other  elements  with  oxygen.  This 
active  principle  is  essential  to  life  and  to  all  other  processes  of 
combustion,  and  when  combined  with  carbon  forms  choke-damp 
or  white  damp;  with  pyrites,  a  common  mineral  in  coal,  sul 
phuretted  hydrogen;  and  with  marsh-gas,  water  or  steam  and 
carbonic  acid.  An  atmosphere  deficient  in  oxygen  will  no  longer 
support  combustion,  and  a  lamp  or  flame  immersed  in  it  is  extin- 
guished; so,  too,  air-breathing  animals  are  suffocated  when  the 
percentage  of  oxygen  in  the  air  is  below  15. 

Nitrogen,  the  predominant  constituent  of  unpolluted  air, 
is  colorless,  odorless,  and  incapable  of  supporting  combustion 
or  animal  life.  It  is  inert  in  its  effect  upon  the  system,  but  extin- 
guishes flame  immersed  in  it. 

Carbonic  Acid,  which  is  not  an  important  constituent  of 
the  atmosphere,  is,  however,  to  be  expected  in  all  abandoned, 
and  unventilated  places.  It  is  given  off  in  quantity  by  the  respira- 
tion of  men  and  animals,  by  the  combustion  of  lamps,  fuel, 
explosives,  timber,  and  organic  matter,  and  by  all  substances 
in  a  state  of  decay  or  fermentation.  It,  also,  is  colorless  and  odor- 


MINE-GASES.  389 

less,  and  at  a  normal  pressure  of  30  inches  of  mercury  and  a 
temperature  of  32°  F.  weighs  128.45  Iks.  Per  thousand  cubic 
feet,  as  against  81  Ibs.  for  an  equal  volume  of  air  under  the  same 
conditions.  It  is  easily  detected,  for,  being  incapable  of  sus- 
taining combustion,  light  burns  dimly  when  immersed  in  it. 
When,  therefore,  the  gas  is  revealed  by  this  action  on  a  flame, 
the  atmosphere  is  at  least  unfit  for  respiration.  The  gas  has  an 
injurious  effect  upon  the  human  system  Air  containing  2  per 
cent  of  it  produces  an  overwhelming  depression  upon  those 
breathing  it;  with  6  per  cent,  lights  are  extinguished;  and  with 
10  per  cent,  it  is  positively  fatal  by  producing  suffocation.  As  it 
accumulates  in  the  lower  part  of  idle  workings  and  where  the 
air  is  stagnant,  no  one  should  venture  into  abandoned  works 
without  having  previously  tested  by  a  flame  the  condition  of 
their  atmosphere,  and  if  the  light  is  extinguished,  the  poison 
must  be  swept  out  by  a  strong  current  of  air,  which,  in  shafts, 
may  be  incited  by  rapidly  raising  and  lowering  a  bundle  of  hay, 
which  is  then  ignited.  Small  volumes  of  carbonic  acid  may 
be  removed  from  the  atmosphere  by  the  use  of  absorbents  like 
lime  or  ammonia.  Its  presence  may  be  chemically  detected  by 
the  milkiness  produced  in  a  test-tube  containing  baryta  hydrate. 

Sulphuretted  Hydrogen,  which  at  30  inches  barometric  pres- 
sure and  a  temperature  of  32°  F.  weighs  94.62  Ibs.  per  thousand 
cubic  feet,  is  an  extremely  poisonous  and  common  gas  occurring 
in  mines.  As  it  is  the  result  of  the  decomposition  of  pyrites, 
always  accompanied  by  heat,  this  gas  is  a  warning  of  incipient 
fires.  In  the  gob  and  abandoned  portions  of  the  mine  the  pyrites 
and  coal-waste  subjected  to  pressure  from  the  roof,  and  in  the 
presence  of  moisture,  decompose  with  the  development  of  heat 
and  possibly  flame,  of  which  the  presence  of  sulphuretted  hydro- 
gen is  an  important  indication.  Being  colorless  but  strongly 
odorous,  its  presence  is  readily  detected.  It  does  not  support 
combustion,  but  is  itself  inflammable.  A  flame  will  burn  in  a 
mixture  of  it  with  air,  but  is  extinguished  in  an  undiluted  atmos- 
phere of  this  noxious  gas,  which  in  this  state  is  fatal  to  life. 

Carbonic  Oxide,  which  is  claimed  to  be  a  normal  gas  exist- 


39°  MANUAL  OF   MINING. 

ent  in  coal-mines,  weighs  78.305  Ibs,  per  one  thousand  cubic 
feet  at  a  temperature  of  32°  F.  and  a  barometric  pressure  of 
30  inches.  This  gas  has  neither  color,  taste,  nor  smell,  and  is 
exceedingly  poisonous.  One  half  per  cent  of  this  gas  in  air  ren- 
ders the  atmosphere  fatal  to  life  if  breathed  for  ten  minutes.  It 
acts  upon  the  system  by  combining  with  the  oxygen  absorbed  in 
the  blood,  forming  a  stable  compound  reducing  the  haemoglobin, 
and  insidiously  and  surely  destroying  the  blood  and  tissues.  As 
carbonic  oxide  is  an  unstable  compound  usually  resulting  from 
the  imperfect  or  secondary  combustion  of  a  gas  or  of  carbon  in 
the  deficiency  of  oxgyen,  the  existence  of  this  gas  in  mines  is 
doubted.  It  is  only  barely  possible  that  it  might  be  found  in 
the  goaf,  where  the  oxidation  of  pyrites  and  the  absorption  of 
oxygen  by  the  fine  coal-dust  may  have  depleted  the  air  of  its 
vital  element,  thereby  giving  rise  to  a  sufficiently  high  temperature 
to  incite  combustion  with  the  development  of  carbonic  oxide. 

Light  Carburetted  Hydrogen  or  Marsh-gas  is  a  stable,  never- 
failing  constituent  among  the  products  of  dry  distillation  of 
organic  matter,  and  exists  as  the  predominant  constituent  of 
the  compound  gas  known  under  the  general  term  of  fire-damp 
because  of  its  ready  ignition  by  flame  with  a  mixture  of  air.  Its 
weight  at  the  normal  barometric  pressure  and  a  temperature  of 
32°  F.  is  about  45.22  Ibs.  per  thousand  cubic  feet.  It  is  absorbed 
in  the  coal,  diffused  through  its  pores,  collected  in  crevices  or 
cavities,  and  even  stored  up  in  reservoirs,  having  been  exuded 
from  the  coal  in  the  early  stages  of  decomposition  of  the  organic 
source  of  the  coal,  or  expressed  during  the  geological  movements 
of  the  earth's  crust.  It  is  not  a  constituent  of  the  coal,  but  is 
entirely  free  from  chemical  combination  with  it,  and  continually 
exudes  with  greater  or  less  violence;  it  is  liberated  in  volumes  by 
falls  in  the  mine- roof,  by  squeezes  or  creeps,  and  by  any  sudden 
fall  in  the  barometric  pressure.  The  deep  portions  of  coal- 
seams  appear  to  be  more  heavily  charged  with  this  compound 
gas  than  are  workings  in  coal-beds  so  close  to  the  surface  as  to 
allow  of  its  escape.  No  coal-seam  should  be  regarded  as  free 
from  liability  to  irruptions  of  this  gas. 


MINE-GASES.  39 1 

The  Occlusion  of  Gas  in  Coal. — Marsh-gas  exists  free  in 
the  pores  of  all  coals,  but  must  not  be  confounded  with  the  gases 
chemically  compounded  with  the  coal.  It  is  under  high  pressure,, 
and  is  given  off  into  the  workings  of  the  mine  with  greater  or  less 
violence  from  the  fissures  or  crevices,  sometimes  without  warn- 
ing, sometimes  accompanied  by  the  heaving  of  the  floor  and  the 
trembling  of  the  roof,  but  always  more  or  less  distinctly  audible. 
The  volume  so  emitted  from  the  coal  varies  in  amount  from  9  to 
100  cu.  ft.  per  ton.  From  anthracite  the  discharge  is  most  copi- 
ous, and  from  bituminous  coal -the  least. 

Frequently  it  escapes  into  the  mine  without  warning;  its 
presence  is  not  always  detected  or  manifested,  as  it  accumulates 
in  a  nook  or  under  a  platform  until  some  untoward  circumstance 
brings  it  into  contact  with  a  naked  light.  The  slight  hiss  accom- 
panying its  exudation  is  hardly  enough  to  be  distinguishable. 
These  "blowers,"  of  all  sizes,  up  to  the  outbursts  that  for  a  time 
overpower  the  ordinary  ventilating  current,  contain  90  per  cent 
of  marsh-gas,  and  may  be  liberated  anywhere  and  at  any  time. 
For  a  long  time  it  has  been  recognized  as  a  constituent  of  the 
gases  entering  mines.  It  escapes  at  various  springs  and  salt- 
mines; it  has  fed  the  sacred  fires  of  Baku  and  the  mud- volcanoes 
of  Bulganak;  it  has  been  found  in  the  Silver  Islet  mine,  in  the 
iron-mines  of  Alsace,  and  in  the  lead-mines  of  Tuscany;  and 
some  years  ago  an  explosive  gas  was  met  with  in  driving  the  lake 
tunnel  at  Chicago.  It  constitutes  from  40  to  90  per  cent  of  the 
natural  gas,  and  is  obtained  among  the  volatile  and  combustible 
constituents  in  the  ultimate  analysis  of  coal.  While  sinking 
shafts  through  porous  strata  fire-damp  has  been  encountered, 
and  precautions  are  therefore  necessary  in  regions  of  natural 
gas,  or  in  formations  immediately  above  the  coal  horizons. 

Outbursts  of  Gas  from  the  Coal. — Measurements  made  by 
the  Royal  Commission  on  Accidents  in  Mines  have  revealed 
the  pressure  of  gas  to  be  frequently  as  great  as  450  Ibs.  per 
square  inch.  Sudden  outbursts  may  therefore  be  expected  when 
escape  is  afforded,  and  these,  according  to  the  treatise  of  M. 
Roberti  Linterman,  preponderate  in  coal-seams  disturbed  by 


392  MANUAL  OF  MINING. 

faults,  foldings,  or  thinnings-out,  and  are  influenced  by  the  dip 
of  the  seam.  They  occur  without  any  premonitory  symptoms 
and  even  in  districts  heretofore  free  from  gas.  They  are  most 
voluminous  in  the  periods  of  pillar-robbing  or  during  the  ex- 
ploration of  virgin  ground.  The  disengagement  of  gases  in- 
creases both  in  intensity  and  frequency  with  the  depth  of  the 
workings  and  with  the  presence  of  permeable  masses  surrounded 
by  rock-masses  so  hard  and  compact  as  to  constitute  effective 
barriers  against  fire-damp.  In  every  district  are  found  one  or 
more  infested  zones  of  gas. 

Owing,  therefore,  to  the  inevitable  occurrence  of  this  in- 
flammable gas  in  all  coal-seams,  and  the  uncertain  quantity 
which  may  be  thrust  into  contact  with  the  flame  of  the  illumi- 
nating-lamp or  of  the  explosive  employed  in  the  mine,  fire-damp 
is  the  dread  enemy  of  coal-miners.  The  amount  of  gas  which 
renders  the  mine  unsafe  cannot  be  stated,  for,  while  an  atmos- 
phere containing  2  per  cent  or  more  is  neither  injurious  to  life 
nor  dangerous  in  mines,  nevertheless  its  presence  to  that  extent 
in  the  air  discharged  from  the  mine  into  our  atmosphere  indicates 
a  probably  excessive  accumulation  in  some  of  the  workings. 
The  Coal  Commission  of  Austria,  in  its  classification  of  mines 
according  to  the  composition  of  the  air  at  the  outlet  of  their  venti- 
lating-shaft,  regards  those  having  more  than  2  per  cent  of  fire- 
damp and  carbonic  acid  in  their  return  air  as  fiery,  and  those 
having  less  than  i  per  cent  of  gas  admixture  as  safe. 

The  Effects  of  Mine-gases  upon  Life  or  Flame. — Mine-gases 
mixed  with  air  do  not  have  the  same  effect  upon  life  or  an  illu- 
minating-flame. In  an  undiluted  state  each  and  all  extinguish 
flame  and  do  not  sustain  life.  When  mixed  with  air  the  carbonic 
oxide  and  sulphuretted  hydrogen  are  poisonous,  while  at  the  same 
time  they  support  combustion.  The  marsh-gas  is  inert  in  its 
influence  upon  life,  but  is  capable  of  ignition;  while  the  carbonic 
acid  is  depressive  in  its  influence  both  upon  flame  and  upon 
life.  There  are  hence  two  methods  of  discovering  the  probable 
unsuitability  of  air  for  respiration — the  extinction  of  flame  by 
an  excess  of  carbonic  acid,  and  the  flame  aureole  from  the  com- 


MINE-GASES.  393 

bustion  of  carbonic  oxide,  the  odor  of  sulphuretted  hydrogen 
being  a  sufficiently  strong  warning  without  other  index  of  its 
noxious  presence.  Fortunately,  however,  the  presence  cf  car- 
bonic oxide  need  only  be  feared  after  a  fire-damp  explosion  or 
after  a  blast  of  one  of  the  lower  grades  of  black  powder. 

Explosive  Gaseous  Mixtures. — Marsh-gas  with  air  will  burn 
freely,  but  when  the  proportion  of  gas  reaches  a  certain  specific 
amount  the  ignition  may  take  place  rapidly,  and  if  the  products 
of  combustion  cannot  escape  equally  rapidly,  explosion  ensues, 
the  force  of  the  explosion  and  the  dangerous  degree  of  dilution 
of  gas  varying  with  the  different  gases. 

The  range  of  proportion  of  gas  dilution  between  the  lower 
and  the  upper  explosive  mixture  is  least  in  the  case  of  fire-damp, 
which  may  vary  in  amount  between  5  and  13  per  cent  of  the 
total  volume  of  air;  is  greatest  in  the  case  of  carbonic  oxide , 
which  will  explode  with  any  mixture  containing  between  13 
and  75  per  cent;  while  with  sulphuretted  hydrogen  the  variation 
lies  between  9  and  28  per  cent  of  gas  in  the  mixture.  The  gases 
are  mentioned  also  in  the  order  of  the  decreasing  danger  of 
explosibility,  the  first  offering  the  greatest  risk.  When  the 
atmosphere  in  which  a  flame  is  immersed  contains  a  percentage 
of  gas  approaching  the  explosive  limit,  the  cap  and  nimbus  become 
large,  and  the  flame  almost  invisible.  The  rapidity  with  which 
the  ignition  is  propagated  depends  upon  the  nearness  of  the 
proportions  of  the  admixture  to  the  figures  given,  and  when 
the  maximum  explosive  ratio  as  indicated  above  is  reached,  the 
propagation  is  instantaneous,  and  the  concussiv  force  of  the 
explosion  is  also  a  maximum;  as  the  percentage  of  gas  recedes 
from  these  ratios  or  increases  beyond  the  limiting  explosive 
proportions,  the  violence  of  the  explosion  decreases.  When 
either  of  the  gases  is  undiluted  with  air,  the  light  placed  in  con- 
tact with  it  is  extinguished.  In  other  words,  a  flame  may  be 
immersed  in  the  workings  filled  with  fire-damp;  and  if  the  line 
of  demarcation  between  the  air  and  the  gases  is  "sharp,"  no 
explosion  will  ensue,  but  the  flame  will  enlarge  and,  after  a  little 
fluttering,  become  extinguished.  This  is  equally  true  of  the 


394  MANUAL  OF  MINING. 

other  gases',  though  the  percentage  of  their  accumulation  in 
mine-workings  is  so  low  and  their  diffusion  so  perfect  that  no 
undiluted  accumulation  of  them  is  likely  to  ensue.  When  igni- 
tion or  explosion  takes  place,  the  products  of  combustion  are 
termed  the  "after-damp." 

But  one  means  is  available  for  the  prevention  of  excessive 
accumulations  of  explosive  gases,  and  this  consists  in  supplying 
a  copious  volume  of  air  and  a  thorough  system  of  distribution 
which  will  dilute  the  noxious  emanations  below  the  danger-line. 
This  will  require  careful  examination  daily  of  all  suspected 
places  and  the  enforcement  of  rigid  discipline.  The  airways 
must  be  ample  in  area  to  allow  the  requisite  volume  of  air 
to  pass  without  producing  a  current  of  high  velocity.  In  mines 
using  the  common  Davy  or  the  Clanny  lamps,  the  maximum 
velocity  admissible  is  300  feet  per  minute.  Where  bonneted 
lamps  are  in  use  the  velocity  may  be  800  feet  without  fear  from 
explosion.  Immunity  from  explosions  is  possible  only  by  ad- 
herence to  these  requirements  and  care  in  the  use  of  blasting 
agents. 

Black  Damp;  After-damp. — This  residual  gas,  after  an 
explosion  or  ignition  of  gas,  will  extinguish  a  flame  because  of 
the  deficiency  of  oxygen,  and  is  theoretically  composed  of  52 
per  cent  of  nitrogen  and  48  per  cent  of  carbonic  acid.  The  per- 
centage of  mixtures  of  oxygen,  nitrogen,  and  carbonic  acid  in 
an  atmosphere  extinctive  of  flame  is  almost  identical  with  that 
of  the  air  expired  from  the  lungs  of  men.  Dr.  F.  Clowes,  as  the 
result  of  a  series  of  experiments  in  relation  to  the  lighting  of 
mines  and  the  behavior  of  lamps,  has  ascertained  that  the  per- 
centage composition  of  the  residual  atmosphere  in  which  flame 
was  extinguished,  is,  oxygen  15.7,  nitrogen  81.1,  carbonic  acid 
3.2,  while  that  of  the  average  exhaled  air  is  oxygen  16.15,  nitro- 
gen 79.9,  and  carbonic  acid  3.95. 

Respiration  in  such  an  atmosphere  is  difficult,  and  produces 
unconsciousness,  followed  by  marked  panting  in  the  effort  to 
supply  oxygen  to  the  lungs.  The  countenance  becomes  swollen 
and  livid,  the  features  distorted,  the  eyes  protrude,  and  as 


MINE-GASES.  395 

asphyxia  is  pronounced,  there  is  a  sudden  cessation  of  the  pul- 
sations of  the  heart  and  of  respiration. 

Usually  some  steam  and  carbonic  oxide  exist  in  the  after- 
damp in  proportions  varying  with  the  temperature  of  explosion 
and  the  initial  proportions  of  air  and  explosive  gas.  In  the 
presence  of  the  latter,  even  in  minute  quantity,  is  the  gravest 
danger.  A  reduction  of  oxygen  to  14  per  cent  or  below  also 
causes  disastrous  results. 

To  the  presence  of  this  carbonic  acid  is  attributed  the  loss  of 
many  -lives  in  a  mine  explosion — more,  in  fact,  than  are  the 
result  of  its  concussion  or  of  contact  with  its  flame.  The  per- 
fectly natural  appearance  of  the  body,  lying  often  by  a  lamp 
still  burning,  proves  the  cause  of  death  to  be  some  insidious 
poison  which  is  combustible,  not  asphyxiation  or  concussion. 
All  members  of  rescuing  parties  entering  the  workings  thereafter 
should  take  due  precautions  against  the  inhalation  of  this  carbonic 
oxide  gas. 

Treatment  for  Asphyxiation. — Persons  overcome  by  any 
gas  may  be  revived  by  blowing  oxygen  into  one  of  the  nostrils, 
the  other  being  closed,  and  by  inducing  artificial  breathing. 
Epsom  salts,  and  water  acidulated  with  vinegar,  are  better  than 
alcoholic  stimulants.  The  warmth  of  the  body  should  be  kept  up, 
and  mustard  plasters  applied  over  the  heart  and  around  the 
ankles.  If  these  produce  no  effect,  recourse  must  be  had  to 
blood-letting  from  the  foot  or  jugular  vein,  and,  as  a  last  resort, 
an  opening  into  the  trachea,  through  which  pure  air  is  forced. 

Those  overcome  by  the  inhalation  of  carbonic  oxide  can 
be  resuscitated  only  by  prompt  action  and  a  copious  supply  of 
pure  oxygen  to  the  lungs. 

The  Force  of  the  Explosion. — Two  volumes  of  marsh-gas 
(CH4)  combine  with  19  vols.  of  air  and  develop  23,550  heat- 
units,  giving  a  temperature  of  6064°  F.  (6525°  absolute)  and 
a  pressure  of  185.3  IDS-  per  square  inch.  If  m  be  the  weight  and 
c  the  specific  heat  of  a  gas,  the  heat  required  to  raise  it  /°  is  ex- 
pressed by  mtc.  To  raise  the  2.75  Ibs.  of  CO2,  14  Ibs.  of  N  and 
2.25  Ibs.  of  water  from  52°  F.  to  /°  F.  require 


39°  MANUAL   OF  MINING. 

2.75  Xo.iyn/  =0.470/5 
14      Xo.i74o/=2.4i8/; 
and 

2.25(160+990)  +  2.25  Xo.2675/  =0.602*+  2587; 
whence 

o.47o/+2.4i8/+o.6o2/+ 2587  =23,550. 

The  force  of  the  explosion  of  two  volumes  of  marsh- gas 
developing  23,550  heat-units  may  be  ascertained  to  be  nearly 
30,000  Ibs.  per  square  foot  by  the  following  analysis: 

Assuming  the  initial  temperature  of  the  marsh-gas  and  the 
mine  air  to  be  62°  F.  or  523°  absolute,  because  it  requires  3.886 
heat-units  to  raise  the  aggregate  products  of  combustion  one 
degree,  the  degrees  to  which  the  final  gaseous  products  of  com- 
bustion will  be  raised  are 

23>55o-  3-886  =6064°  F. 

The  volume  which  these  products  seek  to  occupy  is 
(523+ 6064°)  -5-  523  =12.6  atmospheres. 

12.6X14.7  per  square  inch  equals  26,671  Ibs.  per* square  foot. 

The  Barometric  Relation  of  Explosions. — An  attempt  has 
been  made  to  hypothecate  a  relation  between  the  periods  of  gas 
outbursts  and  the  movements  or  seasons  of  low  barometer,  has 
failed  to  show  any  connection.  A  falling  barometer  has  not 
invariably  been  followed  by  a  heavy  discharge  of  gas,  nor  does 
a  study  of  the  tables  show  its  unfailing  precedence  to  the  evolu- 
tion. While  laying  stress  on  the  acknowledged  fact  that 
December  is  the  worst  month,  there  appears  to  be  no  "off  day" 
for  explosions,  which  are  equally  abundant  on  any  day  of  the 
week.  An  excessively  low  barometer  at  the  sea-level  is  28.3 
inches — a  fall  of  only  6  per  cent  of  the  total  pressure,  and  of 
but  i  per  cent,  or  less,  of  the  pressure  of  the  magazine  gas.  Upon 
the  emissions  from  the  pores  of  the  coal  and  from  goaves  the 
-effect  of  a  barometric  depression  is  noticeable.  But  even  here 


MINE-GASES.  397 

an  acre  of  ground  of  standard  thickness  will,  with  a  barometric 
fall  of  o.i  inch,  exude  only  18  cubic  feet  of  mixture  for  every  25 
yards  of  length  of  face  exposed. 

A  tabulation  of  the  barometric  variations  with  reference 
to  mine  explosions  was  made  in  Westphalia,  during  1896,  with 
the  result  that  of  the  42  explosions  recorded  41  are  attributed 
to  fire-damp  alone,  while  in  one  of  them  coal-dust  participated; 
out  of  the  total  number,  27  happened  when  the  air  pressure 
showed  a  tendency  to  fall  suddenly  or  was  at  its  minimum, 
while  in  the  15  others  the  air  pressure  was  at  maximum,  or  showed 
a  tendency  to  rise.  As  regards  the  places  where  they  occurred, 
the  explosions  are  divided  into  eight  in  exploring  or  preparatory 
workings  in  rock,  26  in  preparatory  workings  in  coal,  and  eight 
in  the  actual  getting  of  coal,  while  the  gas  issued  slowly  in  28 
and  suddenly  in  14  cases. 

The  Diffusion  of  Gases. — It  is  fortunate  that  the  gases  evolved 
from  coal  or  produced  by  the  various  processes  of  decomposition, 
combustion,  and  exhalation  do  not  accumulate  in  separate  layers 
in  the  workings,  with  the  heavier  gas  at  the  bottom  and  the  lighter 
one  near  the  top,  except  in  abandoned  places  where  the  air  is 
allowed  to  stagnate;  but  instead  of  this  even  a  little  circulation 
will  set  up  an  individual  motion  of  the  separate  particles  of  the 
gases,  by  which  they  become  gradually  diffused  throughout  the 
mass  until,  after  sufficient  time  has  elapsed  for  the  purpose, 
they  are  found  intimately  blended,  whatever  may  be  their  relative 
densities.  This  is  not  a  chemical  mixture,  but  a  purely  mechanical 
blending,  depending  upon  the  relative  tensions  of  the  gases. 
The  rapidity  of  this  diffusion  into  atmospheric  air  is  inversely 
proportional  to  the  square  root  of  the  density  of  the  penetrating 
gas.  Marsh-gas,  therefore,  mixes  most  readily  with  the  air, 
carbonic  oxide  not  quite  so  readily;  sulphuretted  hydrogen  less 
so,  and  carbonic  acid  making  with  difficulty  an  intimate  mixture 
with  air.  This  principle  of  diffusion  is  an  exceedingly  valuable 
one  to  the  safety  of  the  mine  and  the  purity  of  its  air,  since  the 
more  ready  the  diffusion  of  the  gas  the  more  easily  will  the 
gas  be  cleared  away.  Thus,  by  the  creation  of  an  air-current 


3^8  MANUAL  OF  MINING. 

throughout  the  workings,  the  gases  are  mixed  with  the  circu- 
lating pure  air,  are  diluted,  and  swept  out  of  the  mine. 

Testing  Mine  Air  for  Gas. — The  condition  of  the  coal  work- 
ings is  usually  examined  and  tested  daily  by  the  fire-boss,  one 
of  whose  duties  consists  in  ascertaining  the  degree  of  saturation 
of  the  mine  air  at  every  place  of  work  before  the  men  are  per- 
mitted to  enter.  The  test  is  made  by  a  candle  or  safety-lamp, 
the  flame  of  which  gives  evidence  of  the  presence  of  an  accumula- 
tion of  combustible  gas.  This  method  requires  a  skilful,  steady 
hand  and  considerable  nerve.  Shading  the  flame  of  a  candle  or 
lamp  with  one  hand,  and  raising  it  upward,  the  fire-boss  watches 
the  behavior  of  the  light.  If  any  inflammable  gas  is  present, 
the  flame  elongates  and  becomes  smoky.  In  this  event  the 
test  ceases,  the  flame  is  lowered,  and  the  fire-boss  withdraws. 
The  face  of  the  coal  or  the  room  showing  these  symptoms  of 
danger  is  then  supplied  with  more  air,  the  employees  being  mean- 
while barred  from  entry  to  the  place. 

The  Height  of  the  Flame  in  Gas. — A  gas  gives  evidence  of  its 
presence  upon  a  flame  immersed  in  it  by  the  elongation  of  the 
flame,  surrounded  by  a  blue  nimbus  or  aureole.  The  more  vola- 
tile the  illuminating-oil  used  in  the  lamp,  the  more  sensitive  is  the 
flame  to  the  presence  of  these  gases.  Thus  naphtha,  benzine, 
and  kerosene,  in  the  order  named,  are  far  more  sensitive  indi- 
cators of  gas  than  13  the  heavy  lard-oil. 

The  Height  of  the  Cap  on  the  light  of  an  ordinary  safety-lamp 
depends  upon  the  percentage  of  gas  to  that  of  air  in  the  mix- 
ture. By  observing  the  height  one  may  determine  the  ap- 
proximate percentage  as  follows:  Divide  the  constant  94,000 
by  the  height  of  the  cap  in  eighths  of  an  inch  and  take  the  cube 
root  of  the  quotient.  Thus,  when  a  blue  cap  is  found  in  a  mix- 
ture under  test  to  be  2  inches  in  height,  then  the  cube  root  of 
94,000  divided  by  16  equals  18.  In  other  words,  there  are  18 
parts  of  air  to  i  of  gas.  If,  in  another  test,  the  length  of  the 
blue  cap  be  only  J  of  an  inch,  then  there  will  be  46  parts  of  air 
to  i  of  gas. 

Testing    Lamps. — Ordinary   safety-lamps   do  not   reveal   the 


MIN&GASES.  399 

presence  of  a  quantity  of  gas  less  than  2  per  cent.  The  Hepple- 
wite-Gray  is  more  sensitive  than  the  unbonneted  Davy  or  Clanny. 
It  burns  benzoline,  and  shows  a  cap  \  inch  high  in  the  presence 
of  i  per  cent  of  CH4.  The  Pieler  spirit-lamp  is  always  a  good 
gas-tester,  which  in  air  containing  \  per  cent  of  combustible  gas 
will  give  evidence  of  it  in  a  cap  i  inch  high.  The  Wolf  safety- 
lamp,  burning  naphtha,  shows  a  very  conspicuous  halo  when 
placed  in  a  mixture  of  fire-damp  and  air;  in  a  mixture  of  \  per 
cent  of  gas  the  flame  is  H  inches  in  height,  and  in  2|  per  cent 
of  gas  the  flame  is  broader,  and  may  even  be  extinguished.  The 
Beard- Mackie  lamp  carries  a  graduated  scale  making  the  height 
of  the  cap  visible.  A  bent  inverted  U  rod  has  platinum  wires 
stretched  across  the  arms,  at  suitable  intervals,  which  become 
luminous  by  contact  with  the  flame.  The  highest  wire  furnishes 
the  guide  to  the  test. 

The  Shaw  Gas-testing  Apparatus  recognizes  the  presence  of 
explosive  gas  and  is  used  for  an  approximate  test  of  mine  air. 
It  meets  with  favor  where  the  exact  analysis  of  the  mine  atmos- 
phere is  not  required.  While  the  apparatus  is  capable  of  demon- 
strating the  quantity  of  explosive  gas  in  the  atmosphere,  it  is  inca- 
pable of  distinguishing  between  them,  and  thus  fails  to  furnish 
any  clew  as  to  the  variety  of  gas  therein  contained.  Of  other 
detectors,  those  depending  upon  the  difference  in  density  of  the 
gases  are  unreliable,  because  changes  of  temperature  will  pro- 
duce similar  results.  Aitken's  indicator  is  ingenious,  but  not 
much  more  reliable.  Its  thermometer  is  coated  with  platinum- 
black  and  plaster  of  Paris,  and  when  exposed  to  fire-damp 
becomes  heated.  If  the  difference  of  temperature  between  it 
and  the  normal  air  always  bore  a  comparable  ratio  to  the  per- 
centage of  fire-damp  contained,  it  would  work  well.  The  special 
forms  of  gas-detectors  do  not  serve  for  illumination. 

The  Amount  of  Air  Required  for  Combustion. — In  attempt- 
ing to  specify  the  amount  of  air  required  for  proper  ventilation 
of  a  mine,  we  are  treading  upon  uncertain  ground.  Within 
close  limits  we  may  ascertain  the  amount  required  for  the  vital 
chemical  purposes  of  horse,  light,  and  man.  A  pound  of  carbon 


400  MANUAL   OF  MINIXG. 

requires  for  complete  combustion  25  Ibs.  of  oxygen,  and  pro- 
duces 3§  Ibs.  of  CO2.  Hence  the  ordinary-sized  mining-candle 
burns  up  n.8  cu.  ft.  of"  air,  and  discharges  3  cu.  ft.  of  CO2» 
Eminent  medical  authorities  state  that  a  man  consumes  about 
i  cu.  ft.  of  air  per  minute,  converting  the  life-giving  principle 
into  2.1  cu.  ft.  of  CO2  per  hour.  The  respiration  of  a  horse 
is  about  13  cu.  ft.  CO2  per  hour.  The  deflagration  of  a  pound 
of  explosive  produces  about  2.6  cu.  ft.  According  to  Angus 
Smith,  two  hewers  using  a  J-lb.  candle  and  12  ounces  of 
powder  produce  25^  cu.  ft.  CO2  in  a  shift. 

The  amounts  of  air  sufficient  to  satisfy  the  conditions  of  com- 
bustion during  the  generation  of  the  respective  amounts  of  COj, 
are  small,  and  if  the  exhalations  were  instantly  removed,  the 
theoretical  chemical  supply  would  suffice.  But  the  air  in  the 
confined  spaces  of  mine-workings  is  somewhat  stagnant,  and 
the  atmosphere  is  further  deteriorated  by  the  exhalations  from 
man  and  beast.  Some  of  these  are  not  easily  detected  chemically,, 
but  are  more  deleterious  than  CO2,  which  is  not  the  sole  test 
of  vitiation. 

The  Amount  of  Air  Required  for  Ventilation. — The  hot, 
noisome  emanations,  the  poisonous  exhalations,  the  unconsumed 
azotic  gases,  and  the  exuding  pent-up  gases  from  the  coal  must 
be  rendered  comparatively  harmless.  This  requires  a  large 
volume  of  air  for  their  dilution  and  renewal.  Pure  dry  air  con- 
tains, by  volume,  21  per  cent  of  oxygen,  O,  and  79  per  cent  of 
nitrogen,  N;  and  every  1000  cu.  ft  of  it,  weighing  nearly  81 
Ibs.,  contains  only  about  18.7  Ibs.  of  the  life-supporting  con- 
stituent, the  remainder  being  matter  inert  in  its  physiological 
effects. 

To  furnish  2§  Ibs.  of  oxygen  for  a  pound  of  carbon  would 
require  142.2  cu.  ft.  of  air  for  combustion  alone.  To  dilute 
the  carbonic  acid  produced  to  a  wholesome  degree,  it  will  require 
2260  cu.  ft.  for  each  pound  of  carbon. 

Judging  by  the  rough  test  afforded  by  the  sense  of  smell,  the 
air  of  a  room  ceases  to  be  wholesome  when  it  contains  more  than 
6  parts  of  CO2  in  10,000.  And  to  preserve  the  lowest  standard 


MINE-GASES,  401 

tolerated  by  sanitarians,  i  in  10,000,  the  supply  will  be  propor- 
tioned as  follows:  59  cu.  ft.  per  hour  per  light;  4585  per  horse; 
9192  per  pound  of  powder;  and  1500  per  man  employed.  Com- 
petent writers  vary  in  this  matter,  and  the  statutes  of  the  various 
States  differ  in  their  requirements  (55  to  300  cu.  ft.  per  man  per 
minute).  But  the  allowance  for  a  mine  cannot  be  based  on  the 
single  per-capita  element,  for  it  will  be  seen  later  that  the  friction 
or  "drag"  of  air,  in  moving  through  headings  and  along  faces 
which  increase  with  the  developments,  diminishes  the  volume  of 
air  actually  allowed  to  move.  Moreover,  the  emission  of  gas 
from  the  strata,  proportional  to  the  area  exposed  and  the  char- 
acter of  the  coal,  constitutes  another  and  constant  source  of  pollu- 
tion. In  all  preparatory  and  prospecting  work  in  virgin  ground 
an  extra  allowance  of  fresh  air  is  necessary.  Cognizance  must 
be  taken  of  this  unfailing  source  to  the  extent  of  an  hourly  allow- 
ance of  0.3  cu.  ft.  of  air  per  square  foot  of  working  face,  in  a  dry,, 
dusty,  fiery  mine.  For  a  non-gaseous  seam  o.i  cu.  ft.  will  suffice. 
Some  property-owners  allow  also  200  cu.  ft.  of  air  per  hour  for 
every  acre  of  goaf.  For  the  eruptions  from  the  magazines  no 
provision  can  be  made  except  vigilance  and  discipline. 

The  Water-gauge  (Fig.  154)  consists  of  a  U  tube  whose  arms 
contain  water  and  are  provided  with  measured  scales  graduated 
to  inches  above  and  below  the  normal  of  the  water  in  the 
columns.  If  a  gauge  be  inserted  through  the  stopping  (Fig. 
155)  separating  the  bottoms  of  two  ventilating-shafts  of  the 
mine,  the  water  will  remain  at  a  normal  level,  if  the  tempera- 
tures and  tensions  of  the  gas  and  air  in  the  two  shafts  be  equal;, 
but  if  by  any  cause  the  tension  or  temperature  be  changed  in 
either  shaft,  the  ensuing  difference  in  pressures  will  be  com- 
municated to  the  connecting-arms  of  the  water-gauge  in  such 
manner  that  the  cool  or  denser  air  will  force  down  the  water  column 
in  the  arm  on  its  side  of  the  stopping  and  elevate  the  water  column 
on  the  opposite  side.  This  difference  in  level,  m,  is  read  on  the  at- 
tached scale  and  represents  the  motive  column,  M,  which  is  capable 
of  producing  motion.  If,  now,  this  excessive  pressure  may  be 
allowed  to  expend  itself  in  producing  motion  through  the 


402 


MANUAL  OF  MINING. 


workings  of  the  mine,  in  circulating  the  air  which  ultimately 
is  discharged  through  the  lighter  column  of  upcast,  the 
level  of  the  water  in  the  gauge  will  fall  slightly  until  equilib- 
rium will  be  established,  when  its  difference  in  level  will  rep- 
resent the  difference  in  pressure,  /»,  at  the  bottom  of  the  two 
shafts  on  either  side  of  the  stopping.  The  excessive  pressure, 
P,  is  expended  in  doing  work  of  two  kinds:  (i)  in  overcoming 


FIG.  154. — Water-gauge.          FlO.  155. — The  Position  of  the  Water-gauge 

on  the  Door. 

the  friction  to  the  passage  of  the  current  of  air  through  the  work- 
ings, from  one  shaft  to  the  other,  and  (2)  in  creating  motion. 
The  latter  work  is  measured  by  the  velocity  of  the  outgoing  cur- 
rent, the  former  is  measured  by  the  height  of  the  water  in  the 
gauge,  Fig.  154,  or  the  manometric  depression,  m.  Its  dif- 
ference in  level  constituting  the  water-gauge  reading,  measures 
the  force  which  is  required  to  drive  the  air  through  the  mine. 
It  measures  the  loss  due  to  friction  or  the  "drag"  of  the  mine. 
Be  the  quantity  of  air  large  or  small,  it  gives  no  measure  of  that 
volume,  but,  paradoxical  though  it  may  seem,  only  of  the  power 
of  the  ventilator. 

The  Mine  Resistance  — The  resistance  of  the  mine  is  a  definite 
quantity,  and  bears  no  relation  to  the  capacity  or  qualities  of  the 
ventilating  appliances  or  methods.  The  water-gauge,  there- 
fore, which  measures  this  resistance  is  a  "function  of  the  mine," 
and  by  it  may  be  determined  the  relative  efficiency  of  the  mine  to 
pass  air  through  its  ways.  The  water-gauge  reading  in  the 


MINE-GASES.  403 

majority  of  mines  varies  between  i  inch  and  3  inches.  Few 
have  a  larger  resistance.  The  mine  with  airways  of  large  cross- 
sectional  area  and  with  a  well-distributed  current  should  have  a 
water-gauge  reading,  or  resistance,  not  exceeding  i  inch  for  each 
hundred  thousand  cubic  feet  circulating  through  it.  The  mine 
having  a  larger  ratio  of  water-gauge  reading  than  this  either  has 
airways  of  insufficient  size  or  is  not  receiving  the  current  prop- 
erly regulated. 

The  Equivalent  Orifice  of  the  Mine.  —  The  resistance  of  the 
mine  to  the  passage  of  an  air-current  is  often  expressed  in  the 
term,  the  equivalent  orifice  of  the  mine.  By  this  is  understood 
the  area  of  a  thin  orifice  which  offers  a  resistance  to  the  passage 
of  a  current  of  the  same  volume,  Q,  equal  to  that  which  is  cir- 
culating through  the  mine.  The  equivalent  orifice  of  the  mine, 
A,  bears  a  certain  ratio  to  the  quantity,  Q,  and  to  the  water- 
gauge  reading,  m,  which  is  variously  expressed  by  different 
authors,  the  general  formula  being 


the  value  for  C  varying  between  0.00037  and  0.00066.    That 
usually  taken  in  calculations  is  0.0004. 

The  equivalent  orifice  of  most  mines  varies  between  10  and 
100  sq.  ft.  Inasmuch  as  water  is  about  833  times  as  dense  as  an 
equal  volume  of  air,  the  column  depressed  in  the  water-gauge 
corresponds  to  a  height  of  833^  inches  of  air  at  62°  F.  and  30 
inches  barometer;  the  head,  M,  of  air  measured  in  feet,  to  which 
the  manometric  depression  is  due,  is 

M  =  6g.4m. 

The  value  for  the  corresponding  differential  pressure,  Pt 
in  pounds  per  square  foot,  is 


EXAMPLE.  —  How  does  the  efficiency  of  a  mine  carrying  178,000  cubic  feet 
of  air  compare  with  another  having  284,800  cubic  feet  per  minute?  The 
water-gauge  readings  are  2.7  inches  and  3.4  inches,  respectively. 

0.0004X17,8000 

a=  -  j=  -  =43.36  square  feet; 
V2.7 


404 


MANUAL   OF  MINING. 


0.0004X284,800     _, 
a=  —   —  -7=  -  =61.78  square  feet. 


Their  relative  efficiencies  are  0.70184:1,  respectively. 

The  following  table  illustrates  the  rate  of  decrease  of  water- 
gauge  in  two  mines,  respectively,  carrying  30,000  and  100,000 
cu.  ft.  of  air  per  minute  for  various  equivalent  mine  orifices. 


Equivalent 

Water  Gauge,  Inches. 

Equivalent 

Water  Gauge,  Inches. 

Orifice. 

Orifice, 

Square 
Feet. 

30,000 
Cubic  Feet 

100,000 

Cubic  Feet 

Square 
Feet. 

30,000 
Cubic  Feet 

100,000 

Cubic  Feet 

Air. 

Air. 

Air. 

Air. 

5 

•56 

25 

0.23 

2-56 

10 

•44 

16.00 

3° 

0.16 

I.78 

15 

.64 

7.11 

40 

I.  00 

18 

•45 

5.00 

5° 



0.64 

20 

•36 

4.00 

100 



o.  16 

REFERENCES. 

Behavior  of  Mine  Gases,  M.  &  M.,  Vol.  XXII,  229;  Air  in  Coal  Mines, 
Prof.  Clowes,  Coll.  Guard.,  Jan.  1896,  222;  Respirability  of  Air  in  which  a 
Candle  Flame  is  Extinguished,  Frank  Clowes,  E.  &  M.  Jour.,  LXI,  515;  Mine 
Ventilation,  W.  J.  Mollison,  M.  &  M.,  Feb.  1904;  Air:  Composition  of  the 
Air  in  Mines,  Coll.  Guard.,  Nov.  8,  1895,  890. 

Gases:  Some  Dangerous  Gases,  Coll.  Guard.,  Jan.  i,  1896,  16;  Compo- 
sition of  a  Mine  Atmosphere  in  Russia,  Trans.  M.  &  M.  E.,  XLV,  Part  V, 
no;  Temperature  and  Humidity  Collieries,  Coll  Guard.,  Oct.  25,  1875,  779 J 
Continuous  Sampling  of  Mine  Air,  Coll.  Guard.,  591;  Mar.  27,  1896;  Some 
Mine  Gases,  A.  Heymann,  Jour  Chem.  &  Met.  Soc.  of  S.  Africa,  Aug.  1903 
and  Sept.  1903;  When  is  Fire  Damp  Dangerous  in  a  Coal  Mine?  Coll.  Guard., 

Oct.    10,    1002. 

Fire-damp  Accumulations,  Coll.  Mgr.,  Mar.  1893,  58;  Fire-damp,  Com- 
position of,  Coll.  Guard.,  Dec.  1896,  1170;  On  Experiments  showing  the 
Pressure  of  Gas  in  the  Solid  Coal,  Lindsay  Wood,  N.  E.  I.,  XXX,  163;  Fire- 
damp Apparatus  for  Experimenting  with,  Coll.  Guard.,  Feb.  14,  1896,  317; 
Detection  of  Small  Quantities  of  Gas  in  Air  Dr.  Clowes,  Collieiy  Mgr.,  Jan. 
18,  1895,  17. 

Indicators  Fire  Damp,  Trans.  M.  &  M.  Eng.,  Vol.  XLV,  Part  V,  106; 
Shaw  Gas  Tester,  Mine  Insp.,  Pa.,  1889,  372;  The  Shaw  Gas  Tester,  M.  & 
M.,  Vol.  XX,  85;  Gas  Detectors,  M.  &  M.,  Vol.  XX,  01. 

How  to  Read  a  Water  Gauge,  M.  &  M.,  Vol.  XXII,  85;  Setting  Pipe- 
valves,  Engineer,  June  i,  1904,  33. 


CHAPTER  XIII. 

METHODS  OF  VENTILATION. 

The  Ventilation  System. — To  obtain  circulation  through  the 
mine,  a  conduit  must  be  furnished  by  which  the  warmer  and 
lighter  air  may  ascend  to  be  supplanted  by  cold  or  compressed 
air  entering  by  a  different  compartment;  and  to  maintain  a 
constant  air-current  throughout  the  workings,  both  inlet  and 
outlet  must  be  afforded  for  the  air  by  means  of  two  separate 
entries  or  by  partitions  in  the  one  shaft. 

The  Ventilation  of  Single  Entries. — Shafts,  in  process  of 
sinking,  or  a  mine  having  but  a  single  entry,  may  discharge 
their  vitiated  air  through  the  wooden  air-tight  box-pipe  pro- 
vided for  the  purpose,  or,  if  there  is  small  liability  of  corrosion, 
through  a  galvanized- iron  pipe,  the  remainder  of  the  entry  furnish- 
ing the  inlet.  Because  of  the  wide  difference  in  the  areas  of 
the  two  airways  so  provided,  the  ventilation  is  not  likely  to  be 
good,  and  it  is  far  better  to  divide  the  main  tunnel  or  shaft  or 
mine-working  into  two  compartments  of  nearly  equal  area,  one 
of  which  will  serve  as  an  outgoing  conduit. 

From  the  fact  that  the  current  in  a  single-entry  mine  is  con- 
tinually interrupted  by  the  other  uses  to  which  the  compartment 
is  put,  and  that  there  is  a  liability  to  injury  of  the  partition,  box, 
or  pipe,  this  plan  is  objectionable  when  a  large  volume  of  air 
is  required,  because  the  safety  of  a  great  number  of  men  is  depend- 
ent upon  this  airway  for  escape.  The  wind,  moreover,  disturbs 
the  ventilating  current;  the  movement  of  cars,  cages,  and  rock 
or  coal  in  chutes  is  also  irregular  in  its  influence  upon  it;  and 
the  unusual  heat  from  underground  steam-pipes,  engines,  etc., 

405 


406  MANUAL  OF  MINING. 

sets  up  counter-currents,  though  any  of  the  causes  mentioned 
may  occasionally  have  a  beneficial  effect.  Thus  a  double  entry 
to  the  mine  becomes  not  only  precautionary,  but  also  impera- 
tive; and  as  the  depth  and  extent  of  workings  increase,  the 
insufficiency  of  a  single  entry  becomes  more  and  more  manifest. 
Even  metalliferous  mines  should  be  provided  with  a  double  entry, 
for  the  numerous  caves  that  have  occurred,  penning  in  dozens 
of  men  without  chance  of  escape  unless  the  rescuers  can  reach 
them  before  suffocation  ensues,  and  the  fires  that  frequently 
cut  off  the  employees  from  the  outlet  and  suffocate  them  before 
extinguishment  is  effected,  are  sufficient  arguments  in  favor  of 
double  entry,  even  if  the  necessities  for  better  air  do  not  appeal 
to  the  operators. 

Ventilation  by  Double  Entries. — All  collieries  have  two  out- 
lets, separated  by  a  safe  distance  of  unbroken  rock.  The  upcast, 
advisably,  should  terminate  in  a  large  chimney,  high  enough 
that  its  draught  be  not  influenced  by  changes  of  wind  or  the 
surrounding  buildings.  The  location  of  the  two  entries,  in 
reference  to  each  other,  varies  within  wide  limits.  One  plan 
consists  in  having  them  near  together,  thus  concentrating  the 
plant.  Both  airways  being  carried  with  the  development,  the 
current  passes  through  to  the  extreme  end  of  one  and  returns  by 
the  other.  Then  as  the  work  deepens,  each  lower  lift  is  connected 
with  the  airways  of  the  upper  lift,  and  receives  ventilation  with 
its  advance.  The  other  plan  is  the  "diagonal  system,"  the 
shafts  being  at  the  extremities  of  the  workings.  While  this  is 
well  enough  for  the  long-wall  method,  the  ventilation  must 
meanwhile  suffer  until  the  connection  has  been  made. 

Two  compartments  in  a  single  entry  may  be  easily  obtained 
in  coal  furnishing  sufficient  rock  from  the  roof  or  from  partings 
by  driving  a  wide  gallery  and  walling  it  up  centrally  with  the 
waste;  but  if  there  is  not  rock  enough  for  this,  two  entries  are 
carried,  with  the  usual  pillar  between  them,  having  connecting 
"throughs"  at  intervals  of  less  than  100  feet,  each  being  closed 
as  fast  as  the  next  one  is  completed.  To  ventilate  that  part 
of  each  entry  between  the  last  connection  of  the  entries  and  its 


METHODS   OF   VENTILATION.  407 

face,  it  is  subdivided  by  a  canvas  brattice  fastened  at  the  near 
side  of  the  "through"  and  leading  up  to  the  work.  On  either 
side  of  this  the  current  flows.  The  faces  may  be  connected  by 
pipes  through  the  door  closing  the  intake  entry  without  inter- 
fering with  the  haulage.  .The  practice  of  relying  upon  diffusion 
to  do  the  work  of  ventilation  is  pernicious.  These  remarks  also 
hold  true  regarding  the  "throughs"  connecting  the  rooms  in 
pillar  and  stall  workings,  where  diffusion  is  usually  relied  upon 
for  the  needful  amount  of  oxygen. 

Planning  Airways. — A  large  number  of  the  coal-mines 
depend  for  their  ventilation  solely  upon  natural  means,  and  this 
may  suffice  in  small  mines.  But  as  the  workings  are  extended, 
the  numerous  connections  which  are  necessary  for  development 
or  convenience  of  handling  the  materials  may  be  planned  to 
serve  also  for  ventilating  ways  without  additional  cost. 

In  planning  the  direction  of  gangways  and  of  rooms  in  coal- 
mines, usually  the  question  of  haulage  is  of  the  first  considera- 
tion, unless  it  be  that  the  "cleats"  are  so  pronounced  as  to  de- 
termine the  direction  of  work.  At  the  same  time  due  attention 
must  be  given  to  the  matter  of  ventilation,  that  the  requisite 
amount  of  air  i>e  given  each  working-room,  and  that  too  many 
men  be  not  dependent  upon  the  same  air-current  circulating 
through  the  mine;  whenever  the  mining  conditions  require  a 
subdivision  of  the  incoming  air-current  into  small  currents,  each 
being  distributed  to  its  own  district  and  group  of  men  and  each 
separately  discharged,  it  becomes  evident  that  the  ventilation 
of  such  gaseous  mines  must  receive  special  attention,  not  only 
as  to  the  direction  in  which  the  airways  are  driven  and  their 
cross-sectional  dimensions,  but  also  as  to  the  means  of  produc- 
ing the  supply  of  air.  In  such  cases  the  fresh  air  should  be 
carried,  if  possible,  to  the  deepest  point  in  the  mine,  whence  an 
ascending  current  may  be  conveyed  through  the  workings  until 
it  is  returned  to  the  surface.  Especially  is  this  advisable  in 
steep  coal-seams  carrying  fire-damp. 

The  ventilation  must  be  so  arranged  that  as  many  inde- 
pendent ventilation  districts  as  possible  be  provided  with  sepa- 


4o8  MANUAL  OF  MINING. 

rate  air-currents;  and  especially  must  each  lift  of  workings  be 
supplied  by  the  shortest  way  with  the  necessary  quantity  of 
fresh  air,  while  within  the  separate  lifts  of  workings  the  air- 
current  must  always  be  ascending — except  in  cases  in  which 
the  descending  air-currents  are  not  used  for  any  further  ventila- 
tion purpose,  or  when,  in  certain  well-ventilated  working  places, 
excessive  thrust  of  the  measures  renders  very  difficult  the  keeping 
up  of  special  return  airways. 

In  metal-mines,  where  the  development  is  of  slower  growth, 
the  rock  hard,  and  a  comparatively  few  men  are  at  work,  the 
amount  of  air  required  is  small,  either  for  inhalation  or  for  the 
dilution  of  the  gases  developed  therein;  hence  a  single  shaft 
with  two  compartments  may  suffice,  the  circulation  being  left 
to  natural  sources.  This,  however,  will  be  inadequate  when 
the  shafts  and  workings  reach  a  depth  of  several  hundred  feet, 
in  which  case  other  means  must  be  employed.  The  use  of  com- 
pressed air  for  drills,  pumps,  etc.,  may  supply  the  deficiency  of 
pure  air  which  natural  ventilation  may  fail  to  furnish,  yet  a  fan, 
exhausting  the  air  from  one  outlet  or  forcing  the  air  into  the 
other,  seems  imperative  with  extensive  workings. 

The  Underground  Temperature. — Below  the  depth  where 
atmospheric  changes  have  no  influence,  the  temperature  of  the 
undisturbed  rock  increases  with  every  increase  in  depth.  The 
depth  at  which  the  temperature  of  the  ground  will  be  found  to  be 
invariable  and  equal  to  the  natural  temperature  of  the  locality 
is  about  50  feet  below  the  surface.  Beyond  this  it  is  an  observed 
fact  that  in  all  artificial  openings  the  temperature  of  the  rocks 
increases  for  at  least  a  moderate  depth,  within  which  the  mine 
operator  is  concerned,  at  the  rate  of  about  i°  F.  for  every  68  feet 
of  depth.  This  increment  is  not  constant  for  all  localities,  nor 
indeed  for  the  same  mine,  but  generally  it  may  be  said  that  as 
we  go  down  the  temperature  of  the  mine  increases  more  or  less 
uniformly.  This  increased  heat  is  often  a  great  drawback  to 
mining,  and  will  ultimately  limit  it  apart  from  the  lesser  mechan- 
ical difficulties.  As  to  what  would  constitute  the  limiting  depth 
to  which  mining  may  be  prosecuted,  it  can  but  be  said  that  at 


METHODS  OF   VENTILATION.  4°9 

present  several  mines,  with  the  exception  of  the  Comstock  and 
those  which  are  in  ore-bearing  districts  feeling  the  effects  of 
solfataric  action,  are  working  at  over  4000  feet.  Regarding  the 
exceptions  stated,  it  is  certain  that  unless  some  means  be  discov- 
ered for  rendering  their  lower  levels  habitable,  the  limit  of  mining 
depth  is  soon  reached.  It  is  stated  that  a  28oo-foot  level  of  the 
Yellow  Jacket  Mines  has  been  abandoned  because  of  the  exces- 
sive temperature,  in  many  rooms  of  which  the  miner  is  com- 
pelled to  return  to  a  cooling  station  after  laboring  only  twenty 
minutes. 

An  interesting  report  bearing  upon  this  question  of  the  rate  of 
increase  of  temperature  with  the  depth  of  subterranean  explora- 
tions, made  by  a  sub-committee  of  the  Royal  Commission  on 
Coal,  reaches  the  following  conclusions:  That  the  limit  of  depth 
to  which  mining  is  possible  depends  upon  human  endurance  of 
high  temperature,  and  upon  the  extent  to  which  it  would  be  possi- 
ble to  reduce  the  temperature  of  the  air  which  comes  in  contact 
with  the  heated  rocks ;  that  there  is  no  limit  caused  by  considera- 
tions of  a  mechanical  nature  as  to  the  size  of  rope  for  hoisting- 
engines,  nor  by  any  consideration  of  the  enhanced  expenditure  for 
shaft  sinking,  for  haulage,  or  for  pumping.  Regarding  the  latter, 
the  experts  testifying  before  them  demonstrated  that  water  is 
seldom,  if  ever,  met  with  in  large  quantities  at  great  depths  in 
mines.  It  therefore  appeared  that  this  increase  in  temperature 
is  the  only  element  needing  consideration  regarding  the  limits  of 
prospective  sinkings  or  workings. 

A  summary  of  the  results  of  temperature  observations  made 
under  the  direction  of  the  British  Commission  Committee  shows 
the  mean  increase  of  temperature  per  foot  to  be  0.01563,  or  i°  F. 
in  64  feet,  the  extremes  being  0.0077  'm  tne  Bootle  water- works 
"bore-holes,  and  0.025  in  the  Carrickfergus  shaft.  At  the  Adel- 
bert  shaft,  Prussia,  observations  five  times  a  month,  in  different 
levels,  for  a  year,  could  deduce  no  regular  law  of  increase;  at  the 
3oth  level,  3200  feet,  the  temperature  was  98°  F. 

Natural  Ventilation. — The  temperature  of  the  air  inside  the 
mine  differs  from  that  outside.  The  mass  which  is  the  warmer 


410  MANUAL  OF  MINING. 

will  rise,  enabling  the  colder  mass  to  fall.  Circulation  is  estab- 
lished so  long  as  this  difference  exists. 
So  that,  if  two  openings  be  made  and 
connected  below,  a  current  will  be  estab- 
lished down  the  lower  and  shorter  open- 
ing in  winter,  and  up  the  same  during 
the  summer,  as  the  arrows  (Fig.  156) 
marked  S  indicate.  In  winter  the 

FIG.  i56.-The  Circulation  of  direction  of  the  current  follows  that  of 
the  Currents  in  Summer  and  the  arrows,  W.  The  amount  of  air  thus 
Winter. 

set  into  circulation  by  the  changes  of  the 

exterior  temperature  will  depend  upon  the  relative  difference  of 
temperature  between  the  mine  and  surface,  and  also  upon  the 
depths  of  the  shafts.  When  these  differences  are  slight  it  is  not 
easy  to  predict  the  direction  which  the  current  will  take.  As,  for 
example,  in  the  fall  and  spring  it  will  fluctuate  from  one  to  the 
other.  When,  however,  these  differences  are  great,  a  current 
will  be  set  up  which  tends  to  continue  in  the  same  direction  so 
long  as  these  differences  remain.  Thus  in  summer  the  current 
will  follow  (Fig.  156)  the  arrows,  5;  in  the  fall  little  or  no  current 
will  be  set  up;  in  the  winter  the  current  will  reverse  and  follow 
the  arrows,  W]  in  the  spring  the  conditions  are  again  nearly 
balanced,  and  little  current  will  flow.  When  the  shaft  attains  a 
depth  of  800  feet,  the  subterranean  air  is  always  hotter  and 
lighter  than  the  surface  air  at  any  season;  and  unless  the  two 
outlets  have  a  great  difference  in  elevation,  an  uninterrupted 
current  will  continue,  without  fear  of  reversal,  down  the  lower 
and  shorter  opening. 

While  this  method  may  be  universally  practised  under 
favorable  conditions  in  metal-mines,  it  is  evident  that  in  col- 
lieries one  danger  arises  from  the  reversal  of  current,  for  at  one 
time  the  current,  following  the  arrows  marked  S,  carries  the  air 
through  the  gangways,  whence  it  is  distributed  among  the  work- 
rooms, to  be  returned  to  the  surface  by  way  of  the  longer  and 
deeper  shaft;  but  during  the  other  season  the  air  may  follow 
the  arrows  marked  W,  thus  entering  the  working  places  first,  and 


METHODS  OF   VENTILATION.  411 

departing  thence  through  the  gangways,  making  its  exit  by  the 
lower  or  shorter  shaft.  If,  now,  there  be  a  number  of  abandoned 
rooms  or  goaves  connected  with  the  working-rooms,  it  is  mani- 
fest that  in  the  latter  season  the  air  must  pass  through  them  first 
before  reaching  the  men  at  work,  and  thus  carry  noxious  gases 
with  the  current  to  spread  calamity  by  explosion  or  fire.  Again, 
the  fact  that  no  air  will  circulate  during  the  vernal  seasons  would 
render  the  provision  for  supplementary  means  of  ventilation  im- 
perative. Air-currents  which  have  served  for  ventilating  pre- 
paratory workings  or  prospecting  drifts  in  the  virgin  seam  never 
should  pass  over  stalls  or  working  places  where  men  are  engaged, 
on  its  way  to  the  air-level. 

The  Tension  of  the  Atmosphere. — The  air  possesses,  in 
common  with  all  other  gases,  in  consequence  of  the  repulsion 
between  its  molecules,  a  tendency  to  expand  into  a  greater  space. 
This  indefinite  expansion,  by  reason  of  which  every  gaseous 
fluid,  not  restricted  by  an  extraneous  force,  continues  to  expand 
to  the  tenuity  of  interstellar  space,  results  in  the  creation  of  an 
air-current  whenever  by  an  increase  of  temperature  or  a  diminu- 
tion of  pressure  the  given  mass  of  air  expands  in  opposition  to  the 
attraction  of  the  earth  and  rises  into  the  upper  strata.  This 
upward  flow  will  continue  so  long  as  the  gas  expands  until  the 
resistance  encountered  by  it  is  equal  to,  or  greater  than,  the 
repulsion  among  its  molecules.  It  is  this  readiness  with  which 
gases  tend  to  adjust  themselves  to  the  varying  conditions  of 
temperature  and  pressure  that  plays  so  important  a  part  in  mine 
ventilation.  The  tension  of  a  gas  increases  with  its  compression, 
and  the  density  of  a  given  mass  of  air  is  proportional  to  its  tension. 

The  Weight  of  Air. — This  may  be  calculated  for  different 
temperatures  and  barometric  pressures  by  the  following  formulae: 

The  volumes,  M,  assumed  by  a  given  weight  of  a  gas  are 
inversely  as  the  corresponding  pressures  per  unit  of  surface, 
provided  the  temperature  remains  constant : 

u:u'::p':p. 
If  the  temperatures  change  while  the  pressures  are  constant, 


412  MANUAL  OF  MINING. 

the  volumes,  reduced  to  absolute  zero  (—  461°  F.),  will  be  found 
to  vary  proportionally. 


/  and  T  being,  respectively,  for  up  and  u'  /,  Fahrenheit  readings. 
The  weight  of  a  cubic  foot  of  air  at  a  temperature  /  and  a 
barometric  pressure  B,  in  inches  of  mercury,  is  obtained  by  the 
following  formula,  and  at  a  temperature  T,  is  W,  expressed  as 
follows: 

1.32535  1*253 


A  table  of  weights  of  air  for  various  temperatures  and  a  con- 
stant pressure,  B,  will  be  found  in  Chapter  IX. 

The  Production  of  Draught.  —  The  energy  in  a  mass  of  air 
compressed  to  a  certain  degree  may  be  measured  by  the  work 
restored  by  it  in  expanding,  and  this  energy  may  be  converted 
into  motion  producing  a  current,  or  it  may  result  in  a  pressure 
when  that  tendency  to  motion  is  resisted,  or  when  the  motion 
is  suddenly  arrested. 

That  portion  of  the  energy  stored  up  in  the  air  which  is 
expended  during  its  expansion  in  dynamic  effect  causes  a 
"wind"  or  "draught,"  the  velocity  of  which  depends  upon 
the  difference  in  tension. 

Whether  the  difference  in  tensions  is  produced  by  change  of 
temperature,  or  of  pressure,  the  velocity  acquired  by  any  gas  is 


in  which  H  is  the  head  due  to  the  difference  between  the  ten- 
sions, or  densities,  of  the  initial  state  of  the  cool  or  compressed 
gas  and  the  final  state  of  the  hot  or  expanded  gas.  Atmos- 
pheric air,  at  a  barometric  pressure  of  29.92  inches,  at  a  tem- 
perature of  32°  F.,  when  flowing  into  a  vacuum,  attains  a  velocity, 
in  feet  per  second,  which  is  equal  to 


METHODS  OF   VENTILATION.  413 

V  =\/2£#  =8.02^  — 

=8.02  JI4'7Xlf  =  10758.9  feet. 

^     O.OOIIO 

The  total  difference  of  pressures  per  square  foot  is  represented 
by  P  in  pounds,  and  the  weight  of  a  cubic  foot  of  the  warmer 
or  attenuated  gas  by  W. 

So,  too,  the  velocity  with  which  compressed  air  or  steam 
escaping  freely  from  a  pipe  or  other  reservoir  of  the  same  may 
be  ascertained,  the  value  to  be  supplied  for  H,  the  head  to  which 
the  velocity  will  be  due  being  equal  to  the  pressure  in  pounds 
per  square  inch  under  which  the  gas  exists,  multiplied  by  144 
and  divided  by  the  weight  in  pounds  per  cubic  foot  of  the  ex- 
hausted fluid. 

The  Motive  Column. — This  is  the  head  producing  motion. 
When,  however,  two  masses  of  air  of  equal  height  but  of 
different  tensions,  p  and  />',  are  exerting  a  pressure  upon  one 
another  through  a  connecting  conduit,  the  resulting  difference 
in  pressure  per  unit  of  area  of  base  measures  the  motive  force, 
in  which  case  P  is  the  total  difference  of  aerostatic  pressure  in 
pounds  per  horizontal  square  foot  of  sectional  area  of  base,  and 
W  the  weight  per  cubic  foot  of  the  rising  column  of  air. 

If,  then,  a  column  of  air  at  /°F.,  D  feet  high,  with  a  base 
of  one  square  foot,  be  heated  to  T°  F.,  its  new  height  would  be 
greater  by  some  quantity,  which  we  may  call  M .  If  two  such 
columns  be  connected,  being  of  the  same  depth  but  of  different 
temperatures,  /  and  T  respectively,  the  latter  column  would  be 
lighter  than  that  at  /°  by  a  quantity  D(w—  W}\  and  so  long  as 
this  difference  in  temperature  is  maintained,  this  difference  of 
pressure,  which  we  may  represent  by  P,  ensues,  by  reason  of 
which  the  hot  column  of  air  would  be  driven  upward,  producing 
a  draught  with  a  velocity,  F,  due  to  the  aerostatic  head,  M.  To 
hold  this  force,  P,  in  equilibrium  would  require  a  resistance 
D(w—  W)  pounds  per  square  foot;  or  the  pressure  of  an  addi- 


414  MANUAL  OF  MINING. 

tional  column  of  warm  air  weighing  W  pounds  per  cubic  foot, 
of  a  height  of  M  feet, 

D(w-W)     P         T-t 

•M-     TT7  TIT  *^ 


w       w     46i -H' 

This  quantity  M  is  known  as  the  motive  column  to  which 
is  due  the  velocity  of  the  flow  of  air,  and  if  no  resistance  is  offered 
to  it,  motion  will  take  place.  It  may  be  represented  by  OT  (Fig. 
156),  which  equalizes  the  pressure  of  the  unequally  heated  columns 
of  air  below  the  level  of  the  line  GO.  • 

When  two  such  shafts  are  of  unequal  depth,  as  at  O  and 
M,  Fig.  156,  and  have  equal  exterior  and  interior  tempera- 
tures, a  rarefaction  of  the  air  in  either  one  of  them  not  affect- 
ing the  other  would  result  in  a  diminished  pressure  upon  the 
bottom,  just  as  is  obtained  by  a  difference  in  temperatures; 
a  rising  current  is  established  therein,  with  a  velocity  depend- 

p 
ent  upon  the  ratio  ^,  in  which  P  is  the  difference  in  the  weights 

of  the  two  shaft  columns  of  a  base  one  horizontal  square  foot 
in  area  and  a  height  D;  and  W  is  the  weight  of  a  cubic  foot  of 
the  rarefied  air. 

For  the  purpose  of  mine  ventilation  there  will  be  required 
a  motive  column  much  larger  than  that  here  obtained,  because 
of  the  enormous  friction  of  the  air  in  rubbing  along  the  rough 
surface  of  the  workings,  turning  sharp  corners,  and  squeezing 
through  small  openings.  The  resistance  due  to  this  cause  amounts 
often  to  as  much  as  90  per  cent  of  the  power.  In  other  words 
only  one  tenth  of  the  theoretical  motive  column  becomes  effec- 
tive in  producing  a  current,  and  the  actual  velocity  of  the  air- 
current,  v,  does  not  exceed  one  third  of  the  theoretical  velocity, 
V ,  due  to  the  head,  If. 

The  principle  upon  which  chimney  draughts  for  boiler  or 
other  heating  apparatus  depends  is  also  similar  to  that  here 
described,  excluding,  of  course,  frictional  allowance.  In  chim- 
neys for  boiler-furnace  draughts  the  fire  burns  best  when  W  is 


METHODS  OF   VENTILATION.  415 

0.530,  and  the  height  of  the  manometric  column  in  the  chim- 
ney is  about  one  half  an  inch  of  water. 

It  is  evident,  therefore,  that  the  height  of  a  motive  column 
depends  upon  the  difference  in  temperatures  or  a  difference  in 
tensions,  or  both,  of  the  gaseous  mixture  contained  in  the  two 
shafts  or  entries  to  the  mine.  A  measure  for  this  motive  column 
may  be  had  in  feet  of  head  of  pressure  per  square  foot  of  area  of 
the  base,  or  in  the  number  of  inches  of  a  water  column  in  the 
manometer  which  corresponds  to  this  weight.  Insomuch  as  a 
column  of  water  i  inch  high  and  a  square  foot  in  area  of  base 
weighs  5.184  Ibs.,  the  height,  m,  of  a  water-gauge  column  which 
will  balance  the  pressure  P  is  equal  to  P-v-5.i84. 

Let  M  be  the  head  corresponding  to   the   motive    column, 

v* 
v  the  velocity  of  flow  of   the  upcast  air  per  second;   then  is  — 

the  effective  velocity  head  of  the  issuing  air;  and  if  W  is  the 
weight  of  a  cubic  foot  of  the  warm  or  attenuated  rising  air,  the 
theoretical  energy  of  the  moving  air  per  second  is  WM  ,  and  the 

v2 
effective  or  actual  energy  is  W  —  ,  or  0.015  53  WV8.     That  por- 

tion of  the  energy  which  is  consumed  in  overcoming  the  friction 
of  the  mine  is  therefore  W(M  —  0.015  53^2).  It  is  this  lost  energy 
which  is  measured  by  the  water-gauge.  As  the  mine  resistances 
are  reduced,  so  the  water-gauge  reading  is  reduced,  and  the 
efficiency  of  the  mine  increases,  permitting  a  greater  actual 
return  from  the  expenditure  of  the  same  potential  force. 

The  height,  w,  of  the  water  column  being  measured  in  inches, 
the  number  of  horse-powers,  H,  necessary  to  produce  a  ven- 
tilating current  of  Q  cubic  feet  per  minute  is  known  by  the  fol- 
lowin formula  : 


The  indication,  therefore,  which  the  water-gauge  reading 
gives  of  the  ventilating  force  is  evident  in  the  above  formula  — 
that  for  a  given  quantity  of  air,  Q,  in  circulation,  the  horse-power 
necessary  to  produce  ventilation  increases  with  the  resistance  of 
the  mine. 


41 6  MANUAL  OF  MINING. 

The  Systems  of  Producing  a  Ventilating  Current. — The  sev- 
eral methods  of  accelerating  the  natural  ventilation  and  dis- 
tributing air  properly  throughout  the  mine  contemplate  some 
system  either  of  decreasing  the  tension  of  the  mine  air  to  enable 
the  return  current  to  ascend  to  the  surface,  or  of  increasing  its 
tension  by  the  use  of  a  compressor  to  force  atmospheric  air  into 
the  mine.  The  several  means  by  which  these  results  are  attained 
may  be,  first,  a  furnace  built  at  the  bottom  of  the  outlet  shaft, 
or  a  fan;  or,  second,  a  blowing,  propelling,  or  air-compressing 
fan  at  the  mouth  of  the  inlet  shaft.  By  either  of  these  methods 
a  different  state  of  tension  is  produced  in  the  two  shafts  con- 
nected below,  and,  in  the  effort  to  establish  equilibrium,  the  air  is 
set  in  motion,  a  draught  is  created,  and  a  current  is  established 
that  flows  through  the  air-courses  at  a  velocity  dependent  upon 
the  head  due  to  the  difference  in  pressures,  as  has  been  seen  on 
the  preceding  page. 

Furnaces  are  employed  for  increasing  the  temperature  and 
are  constructed  in  such  manner  as  to  be  remote  from  direct 
contact  with  the  coal,  yet  in  close  proximity  to  the  shaft  which 
constitutes  the  outlet  for  the  mine  air,  and  in  a  gallery  through 
which  circulates  air  from  the  workings.  The  pit  selected  for 
the  outlet  should  be  that  one  which  would  naturally  carry  the 
flow  in  winter.  The  furnace  is  simply  a  fireplace,  walled  and 
roofed  by  a  fire-brick  or  common-brick  arch  (Figs.  157  and  158). 
When  special  care  is  taken,  a  second  wall  is  built  outside  and 
over,  with  an  air-space  between,  to  isolate  it  from  the  coal  and 
prevent  fire.  If  the  roof  is  wet,  a  double  arch  must  surmount 
the  furnace,  as  otherwise  the  steam  generated  will  burst  the  arch. 
If  the  mine  is  fiery,  or  considerable  dust  is  floating,  care  must 
be  taken  that  the  gases  are  well  diffused,  or  else  the  current  must 
not  be  brought  into  close  proximity  with  the  fire.  In  such  cases 
the  current  is  split,  a  small  portion  being  heated  over  the  .fire, 
the  remainder  passing  through  a  "dumb-channel,"  entering  the 
upcast  50  feet  or  so  above.  A  still  safer  plan  passes  all  of  the 
fiery  current  through  the  channel,  and  feeds  the  furnace  by  a 
split  current  of  fresh  air  direct  from  the  intake. 


METHODS  OF   VENTILATION. 


417 


The  size  of  the  grate  depends  upon  the  work  to  be  done. 
Its  bars  are  3  feet  from  the  floor,  slanting  upward  toward  the 


FIG.  157. — Longitudinal  Section  of  a  Furnace. 

shaft  i  to  6,  distance  to  the  roof  4  or  5  feet.  The  width,  wall  to 
wall,  is  6  feet  and  its  length  from  4  to  12  feet,  according  to  the 
volume  of  air  to  be  moved,  which  is  about  1500  cu.  ft.  per  square 
foot  of  fire  surface  on  a  properly  constructed  furnace. 


FIG.  158. — Cross-section  of  a  Furnace. 

An  ordinary  furnace  of  34  sq.  ft.  heating-surface,  costing 
$130,  will  heat  a  column  of  air  such  as  will  furnish  29,000  cu.  ft. 
per  minute.  A  large  number  of  furnaces  10X12  furnish  200,000 
cu.  ft.  The  cross-sectional  area  must  be  50  per  cent  greater 
than  the  upcast  airway,  and  the  shape  capable  of  regulation  by 


41 8  MANUAL  OF' MINING. 

double-sliding  iron  doors,  to  produce  varying  degrees  of  con- 
traction and  of  combustion.  The  fire  is  spread  over  its  entire 
width,  and  over  only  as  much  of  its  length  as  is  necessary  to  fur- 
nish an  adequate  motive  column,  at  a  temperature  of  140°  to 
160°  F.  Emergencies,  as  low  barometer  and  high  thermometer, 
and  the  cleaning  of  the  grates,  require  other  and  more  heating- 
surface.  The  coal  consumed  is  2  to  5  tons  per  day,  spread  thin 
and  evenly  over  the  bars,  and  fed  from  both  ends,  on  a  long  fur- 
nace. This  rate  is  40  to  70  Ibs.  per  hourly  H.P.  of  work  done 
on  the  air.  Attendance,  etc.,  is  $5  per  day. 

Q  being  the  quantity  of  air  in  cubic  feet  per  minute,  W  the 
weight  of  a  cubic  foot  of  return  air,  T  being  the  temperature, 
F.°,  of  the  up-cast  air,  and  t  that  of  the  air  in  the  return  airway, 
the  number  of  pounds  of  coal  consumed  by  the  furnace  per  hour 
is 

x= 

The  area,  F,  of  the  grate-surface  in  square  feet  is  about  one 
tenth  of  the  hourly  coal  consumption,  in  pounds,  and  its  rela- 
tion to  the  depth,  D,  of  the  furnace  below  the  surface  is  known 
by  the  expression 

FQP=  i,7i6,ooo\/Z>, 

P  being  the  manometric  depression  in  pounds  per  square  foot, 
and  Q  the  volume  of  air  per  minute. 

The  Limits  of  Furnace  Ventilation. — A  method  so  simple 
and  cheap  in  construction  and  easy  of  management  presents 
advantages  which  have  long  commended  it  to  mine  operators; 
nevertheless  the  difficulties  with  its  use,  the  dangers  which  attend 
the  exposing  of  an  open  fire  in  gaseous  districts  without  the 
possibility  of  introducing  a  safeguard,  the  numerous  calamities 
traced  to  the  furnace  which  has  fired  either  the  solid  coal  sur- 
rounding it,  the  gases  in  the  return  air,  the  timbers  of  the  shaft, 
or  even  the  surface  plant,  and  its  lack  of  economy  in  shallow 
pits,  were  soon  made  manifest.  The  atmospheric  changes  of 
the  seasons  reduce  its  efficiency,  a  decrease  in  the  barometric 


METHODS   OF   VENTILATION.  419 

pressure  and  an  increase  in  the  surface  atmosphere  reduced 
the  action  of  the  furnace,  and  notwithstanding  its  great  superi- 
ority over  many  other  mechanical  appliances,  it  has  gradually 
been  supplanted  by  fans.  The  power  of  the  furnace  increases 
arithmetically  with  the  temperature,  and  that  with  the  amount 
of  fuel  burned.  The  quantity  of  coal  that  can  be  consumed 
upon  a  given  area  is  limited.  The  resistance  of  the  mine  increases 
on  the  other  hand  geometrically  with  the  square  of  the  velocity 
of  the  current,  and  it  is  therefore  manifest  that  between  the 
several  conditions  the  furnace  limit  is  soon  reached.  Many 
furnaces  may  be  cited  supplying  to  the  mine  over  20x3,000  cu.  ft. 
of  air  per  minute;  and  enormous  as  they  are,  their  cost  is  very 
little  less  than  that  of  a  modern  fan  of  large  size;  but  when  we 
contemplate  the  huge  pile  of  coal  thus  consumed  for  the  produc- 
tion of  the  current,  we  are  forced  to  the  conclusion  that  efficient 
furnace  ventilation  is  a  luxury  which  the  coal  trade  cannot  long 
endure.  Perhaps  as  the  depth  of  the  collieries  increases  to 
about  2000  feet  the  furnace  may  be  reinstated. 

With  the  atmospheric  air  at  62°  F.,  and  the  furnace-heated 
air  at  132°  F.,  the  water-gauge  depressions,  m,  produced  at  various 
depths  of  furnace  are  as  follows: 


D. 

m. 

D. 

m. 

50  feet, 

0.086  inch. 

1000  feet, 

1-735 

400  " 

0.694    " 

2OOO      " 

3-471 

700  " 

1.215     " 

4000   " 

6-943 

These  are  in  accordance  with  the  formula  for  estimating  the 
manometric  depression: 

W  I    T-t  \ 

- 


EXAMPLE.  —  A  colliery  has  two  shafts  1000  feet  deep,  12  feet  in  diameter; 
temperature  in  the  downcast  is  60°  F.;  barometric  pressure  is  30  inches. 
150,000  cubic  feet  of  air  are  supplied  per  minute  by  a  furnace.  Required 
the  temperature  of  the  upcast  and  the  horse-power  necessary  to  produce  the 
ventilation,  the  mine  being  supposed  to  show  a  water-gauge  resistance  of 
2  inches  of  water.  520°  F.  and  163.5  H.P. 


420  MANUAL   OF  MINING. 

Assume  the  coefficient  of  friction  for  the  smooth  shafts  to  be  as  great  as 
that  of  the  rough  mine  galleries;    then 

O.OOOOOOO2I7X  1  000X37.  7X(i5o,ooo)2 
/>=-  (113)'  =I2'8- 

Each  shaft  therefore  offers  a  resistance  equal  to  12.8  Ibs.  per  square  foot. 
The  total  resistance  then  is  25.6+10.368=35.968.     The  work  done  in  ven- 
tilating is  150,000X35,968=5,395,500  ft.-lbs.,  or  163.5  horse-power. 
The  temperature  of  the  upcast  shaft  is 

T-6o 


T  .  ---         -  . 
39-759  521 

Or,  by  another  method:  A  cubic  foot  of  air  at  60°  F.  and  30°  barometer  weighs 
0.0766  Ib.  150,000  cubic  feet  of  the  circulating  air  weigh  11,490  Ibs.  Since 
the  furnace  is  performing  5,395,500  ft.-lbs.  of  work  upon  11,490  Ibs.  of  air, 
the  height  through  which  it  is  moved  is  470  feet  in  one  minute.  Then  M  is 
470  ft.,  /=6o°F.,  and  D=iooo  ft. 


From  this  it  is  seen  that  the  temperature  of  the  upcast  air  necessary  to  force 
150,000  cubic  feet  of  air  through  the  mine  is  dangerously  high.  The  furnace 
must  be  replaced  by  the  exhaust-fan,  or  the  frictional  resistances  must  be 
reduced  by  enlarging  the  entry-ways. 

What  should  be  the  size  of  the  airway  shafts  in  the  above  case,  that  the 
upcast  air  be  not  hotter  than  190°  F.  ?  By  substitution  above,  M  is  found  to 
be  200  ft.;  the  work  is  then  11,490X200=2,300,000  ft.-lbs.  (70  H.P.).  This 
value  requires  that  p  should  not  exceed  15.3  Ibs.,  which  limits  the  shaft's 
resistances  to  4.932  Ibs.,  or  2.466  Ibs.  each.  In  order  to  obtain  so  low  a 
friction  the  areas  are  enlarged  to  a  radius  of  16.66  feet. 

pa*=jlmq\     or    p(xr>y=jl(2xr}q\ 

Types  of  Fans.  —  Of  the  various  mechanical  ventilators  tried 
in  mines,  fans  remain  our  main  reliance  at  the  present  time. 
As  the  furnace  has  in  the  past  supplanted  various  mechanical 
devices  in  the  form  of  pumps  and  trompes,  so  fans  built  on  various 
principles  have  succeeded  the  furnace  and  the  steam-jet.  There 
are  two  classes  or  types  of  fans:  (i)  blowers,  either  rotary  or 
reciprocating,  and  (2)  fans,  propeller  or  centrifugal.  Those 
of  one  type  sweep  out  a  fixed  volume  of  air  at  each  revolution 


METHODS   OF   VENTILATION.  42* 

and  are  known  as  the  definite-volume  exhausters,  under  which, 
head  come  the  Root,  Baker,  Lemielle,  Cooke,  and  Fabry.  In. 
the  other  class,  acting  centrifugally  upon  the  air,  we  have  a 
simple  revolving  wheel  always  working  in  one  direction,  pro- 
ducing by  its  rotation  a  pressure  or  a  rarefaction  the  degree  of 
which  depends  upon  its  speed.  Of  these  we  have  the  Guibal, 
Waddle,  Walker,  and  Schiele  ventilators. 

The  Trompe  is  a  simple  application  of  the  injector  princi- 
ple,— water  falling  in  the  cylinder  and  carrying  with  it  air,  creates 
a  small  intake  draught.  The  volume  of  air,  compared  with  the 
quantity  of  water  used,  is  so  insignificant  that,  unless  an  espe- 
cially favorable  means  be  provided  for  carrying  off  the  water,  the 
ventilation  is  too  expensive  to  be  continued  except  as  a  temporary 
expedient. 

Pressure  Blowers,  either  rotary  or  reciprocating  in  their 
action,  are  of  general  use  in  America,  being  represented  by  the 
Root,  Baker,  and  Champion  on  the  one  hand,  and  air-compressor 
and  other  reciprocators  on  the  other.  The  blower  forces  the 
air  through  the  intake  compartment  of  the  mine,  which  dis- 
charges it  at  the  upcast.  These  blowers  or  force-fans  are  much 
in  vogue  for  small  workings  and  as  expedients  in  furnishing 
a  separate  ventilation  for  stopes  and  drifts;  but  few  are  employed 
in  coal-mines  to  produce  the  total  ventilation  there  required. 
In  metal-mines,  however,  they  are  largely  depended  upon,  though 
they  supply  a  pressure  higher  than  that  ordinarily  required  to 
overcome  the  resistance  of  the  mine.  They  produce  air  by 
reason  of  their  high  speed  at  a  pressure  often  attaining  10  Ibs. 
per  square  inch,  whereas  a  mine  requires  an  initial  pressure  only 
sufficient  to  overcome  its  resistance,  which  is  rarely  greater  than 
10  Ibs.  per  square  foot  of  base.  The  blower  is  a  small  radial 
wheel  revolving  freely  in  a  casing  and  nearly  touching  its  sides. 
By  a  central  opening  on  either  side  the  air  is  admitted  to  te 
acted  on  and  set  into  rotary  motion.  These  blowers  may  be 
had  in  sizes  capable  of  furnishing  as  much  as  16,000  cu.  ft.  per 
minute,  requiring  from  i  to  15  horse-power  for  their  operation. 
Some  of  the  blowers  are  capable  of  a  ready  alteration  from  a 


422  MANUAL  OF  MINING. 

t>lower  to  an  exhaust,  or  the  reverse,  which  fact  recommends 
them  particularly  for  wide  shafts  which  are  liable  to  freeze  during 
winter.  This  is  particularly  advantageous  in  metal-mines, 
^where  it  makes  very  little  difference  which  way  the  current  moves. 
En  collieries,  however,  as  has  been  seen,  this  is  not  feasible. 

The  Root  Blower  or  force-fan  consists  of  two  interlocking 
timpellers  revolving  side  by  side  in  very  close  connection,  without 
actually  touching  one  another  or  the  enclosing  case.  They  are 
•made  of  cast  iron  accurately  bored  and  dressed  to  a  true  surface, 
•so  that,  while  practically  no  air  escapes,  there  is  also  no  internal 
wear.  At  each  revolution  a  definite  volume  of  air  enters,  is 
enclosed,  and  discharged  either  at  the  top,  the  bottom,  or  the 
side.  They  are  driven  by  a  pair  of  external  gears  at  a  speed 
ordinarily  of  from  250  to  500  revolutions  per  minute.  The 
extremities  of  the  revolving  arms  of  the  impeller  section  are  of 
an  acorn  shape,  or  their  surfaces  are  arcs  either  of  true  circles 
.  or  of  cycloids. 

The  Fabry  Blower,  which  resembles  the  Root,  is  much  used 
in  the  north  of  France  and  Belgium.  Two  fans,  each  having 
three  broad  blades  arranged  radially,  are  hung  in  a  chamber. 
They  revolve  with  equal  velocities  in  opposite  directions,  the 
blades  coming  in  contact,  isolating  a  quantity  of  air,  and  expel- 
ling it  into  the  atmosphere.  The  success  of  this  blower  is  attrib- 
uted to  the  fact  that  there  are  no  joints  in  it. 

The  Baker  Rotary  Force-fan  has  inside  of  its  casing  three 
-drums,  each  being  an  independent  casting  turning  truly  and 
balanced  perfectly  to  insure  a  steady  motion.  The  upper  drum, 
which  receives  the  power  from  the  engine,  does  all  the  work 
of  blowing,  while  the  two  lower  drums  serve  as  valves  to  prevent 
the  air  from  escaping. 

Cooke's  Fan  is  a  positive  machine.  An  eccentric  drum 
revolves  inside  of  a  1 2-foot  circular  case,  very  close  to  which  is 
held  a  swinging  shutter  that  cuts  off  the  entering  current  from 
the  discharge.  The  inlet  and  outlet  portion  occupies  235°  of  a 
-revolution.  At  the  Lofthouse  iron-mines  are  seen  two  of  these 
side  by  side,  the  drums  being  placed  opposite  each  other  on  the 


METHODS   OF   VENTILATION.  42$ 

shaft,    so   that   the    revolving    mass   is   balanced,    the  discharge 
equalized,  and  the  efficiency  raised. 

The  Lemielle  Blower  is  a  species  of  air-pump,  complicated! 
and  leaky,  producing  large  volumes  under  great  rarefaction. 
It  consists  of  a  vertical  cylinder,  within  which  a  second  revolves, 
eccentrically;  on  this  latter  are  two  or  more  vanes,  which  in* 
one  part  of  the  revolution  lie  close  to  the  shutter,  and  in  another 
open  and  expel  the  air. 

The  reciprocating  blowers  have  been  displaced  almost, 
entirely  by  the  rotary  blowers,  either  class  being  capable  of  a 
reversal  of  rotation  to  force  air  into  or  exhaust  air  from  the  mine, 
as  desired.  The  power  required  to  drive  the  force-fan  depends- 
upon  the  volume  and  pressure  of  air  exhausted  or  discharged;, 
but  the  rule  usually  followed  for  computing  the  net  power  in  a 
given  volume  at  different  pressures  is  'to  multiply  the  number  of 
cubic  feet  delivered  per  minute  by  the  pressure  in  pounds  per 
square  foot  at  the  blower,  and  the  product  by  0.00003;  the  quo- 
tient will  give  the  net  horse-power  required  to  drive  the  fan. 

The  Centrifugal  Fans,  which  are  used  almost  exclusively 
in  America,  may  be  divided  into  two  great  classes:  (i)  Those 
which  are  called  open-running,  by  which  we  mean  that  they 
are  free  and  discharge  their  air  all  around  the  circumference; 
and  (2)  those  called  closed- running  fans,  which  have  but  a  re- 
stricted opening  for  the  discharge  of  the  air.  Those  of  either 
class  are  made  large  in  diameter  and  are  driven  at  a  relatively 
small  angular  velocity,  though  a  few,  such  as  the  Schiele,  are  of 
small  diameter,  running  at  a  high  angular  velocity.  They 
produce  large  volumes  of  air  at  a  low  pressure,  and  may  be- 
reversed  in  motion  to  exhaust  or  to  force  air.  The  diameter 
of  the  fans  of  this  class  may  be  and  is  occasionally  as  high. 
as  50  feet,  those  of  small  diameter  being  regarded  as  unnec- 
essarily cumbrous.  The  action  of  all  fans  is  based  upon  the 
general  law  that  bodies  in  motion  tend  to  travel  in  straight  lines,, 
resisting  any  attempt  at  diversion  from  this  path,  in  consequence 
of  which,  when  the  fan  is  set  in  motion,  its  blades  come  in  con- 
tact with  its  interior  air,  the  particles  of  which  are  at  rest  and 


424  MANUAL  OF  MINING. 

resist  rotation.  When,  however,  the  particles  do  move,  their 
endeavor  to  travel  in  straight  lines  results  in  their  making  for 
the  circumference,  producing  thereby  in  the  central  portion  of 
the  fan  a  partial  vacuum,  which  is  replaced  by  the  air  external  to 
the  fan.  So  long  as  the  rotation  of  the  blades  continues,  so  long 
will  this  current  be  produced  and  maintained,  the  pressure  of 
which  will  increase  as  the  peripheral  speed  increases. 

Open-running  Fans. — These  fans  include  the  Waddle,  Biram, 
Naysmith,  and  Hopton  patterns,  all  of  which  are  essentially 
similar  to  the  first  named.  The  Waddle  is  a  self-contained  fan, 
in  that  there  is  no  fixed  casing,  and  the  whole  machine  re- 
volves. Its  form  is  practically  that  of  a  light  hollow  disc  of 
wrought  iron,  the  blades  and  casing  being  wholly  riveted  together. 
The  air  enters  by  a  straight  lead  at  one  side  only,  and  passes 
through  curved  and  gradually  narrowing  channels  to  the  cir- 
cumference, the  blades  being  bent  at  first  to  incline  slightly  back- 
wards, the  alternate  blades  extending  not  more  than  one  half  the 
distance  between  the  circumference  and  the  inlet.  The  passages, 
by  their  contraction,  are  so  made  that  the  circumference  at  any 
point  multiplied  by  the  cross-sectional  area  at  that  point  is  a 
constant  quantity.  The  outer  circumference  of  the  fan  is  bell- 
mouthed. 

A  fan  of  9  feet  diameter  circulates  80,000  cu.  ft.  with  a  water- 
gauge  of  2  inches.  One  of  45  feet,  driven  by  an  engine  with 
40"  X42"  cylinder,  at  a  boiler  pressure  of  80  Ibs.  per  square  inch, 
has  given  a  volume  of  over  550,000  cu.  ft.  at  42  revolutions. 

The  Hopton  has  an  inlet  on  each  side  of  the  central  diaphragm 
with  backward-curving  blades,  and  a  construction  very  simple. 
The  revolving  portion  consists  of  the  arms  and  blades  working 
between  two  brick  walls. 

The  open-running  fans  must,  in  order  to  be  efficient,  dis- 
charge their  air  at  a'  very  low  velocity,  because  the  energy  of 
bodies  in  motion  increases  as  the  square  of  the  velocity,  and 
that  passed  by  the  discharged  air  is,  therefore,  so  much  use- 
less work.  It  is  for  this  reason  that  the  passages  in  Jthe  more 
correct  open-running  fans,  like  that  of  the  Waddle,  are  curved 


METHODS  OF   VENTILATION.  425 

backward.  The  theoretical  depression  which  can  be  produced 
in  fans  of  this  type  is  equal  to  the  height  due  to  its  peripheral 
speed,  Tt  in  feet  per  second. 


The  Closed-running  Fans  are  essentially  of  a  more  massive 
structure  than  those  of  the  open-running  type,  being  of  con- 
siderable width  as  well  as  of  diameter.  Of  this  class  of  fans 
the  Guibal  is  a  type,  the  Schiele  and  the  Walker  Indestructible 
being  similar  in  construction.  Inside  of  a  fire-proof  housing 
a  horizontal  shaft  is  revolved  by  an  engine  or  dynamo,  carrying 
with  it  an  hexagonal  or  square  frame,  on  which  are  built  six  or 
eight  blades.  The  blades  are  flat  and  slightly  curved  at  their 
tips,  sometimes  radially,  and  often  inclined  backwards.  The 
clearance  between  the  tips  of  the  blades  and  the  casing  is  made 
as  little  as  possible,  except  for  a  certain  distance  at  the  bottom, 
through  which  the  air  is  discharged,  the  amount  of  that  opening 
being  regulated  by  an  adjustable  shutter  in  a  gradually  enlarging 
chimney.  The  air  enters  at  the  centre,  whence  it  passes  into  one 
of  the  intervals  between  the  consecutive  blades,  which  form  an 
evasee  canal,  the  speed  of  exit  being  less  than  the  speed  of  entry 

(Fig.  159)- 

In  the  Schiele  fan  the  blades  are  contracted  in  width  from 
inlet  to  outlet,  the  fan  being  surrounded  by  the  usual  spiral  cas- 
ing, into  which  the  air  discharges  all  around  the  circumference, 
the  space  continually  increasing  until  it  reaches  the  chimney. 
The  blades  of  the  Rateau  fan  extend  to  the  centre  of  the  fan,  and 
have  a  peculiar  curvature  slightly  forward,  and  also  a  curvature 
in  the  line  of  the  fan-shaft.  Immediately  in  front  of  them  is  a 
cone  terminating  in  a  point.  The  Capell  fan,  of  equal  power 
with  the  Guibal,  is  smaller,  and  runs  at  a  higher  speed.  It  has 
two  concentric  shells  besides  its  outer  casing,  in  each  of  which 
are  curved  blades  with  the  convex  side  forward.  The  air  enters 
the  inner  shell,  is  forced  out  through  ports  into  the  second  outer 
shell,  where  it  strikes  the  concave  face  of  the  outer  blade,  and 


426 


MANUAL   OF  MINING. 


thence  is  discharged  at  a  low  velocity  through  the  usual  expand- 
ing exhaust-flue. 

The  Champion  fan,  which  is  really  two  fans  joined  together 
by  a  common  centre  ring,  is  designed  to  propel  the  air  with  a 
minimum  resistance,  the  blades  having  a  backward  curvature. 
The  us-?  of  the  inner  casing  or  hood  and  attendant  diaphragm, 


FIG.  150. — Section  of  a  Fan. 

which  are  hung  on  frames,  renders  it  possible  to  change  the  cur- 
rent at  will,  blowing  to  exhaust,  by  revolving  the  hood  around 
the  fan  without  stopping  the  latter. 

The  Evase"e  Chimney.— The  depression  produced  by  a  cov- 
ered ventilator  with  an  expanding  chimney  is  twice  that  of  the 
uncovered  or  open-running  type,  and  is  double  the  height,  due 
to  the  tangential  speed  of  its  blade-tips.  The  use  of  the  chimney 
gives  to  this  type  of  fan  the  enormous  advantage  over  the  other 
that  the  air  may  be  discharged  from  the  fan  at  a  higher  velocity 
without  any  material  loss  of  energy.  The  gradually  increasing 
space  into  which  they  discharge  reduces  the  velocity  and  utilizes 
all  the  energy  in  giving  motion  to  the  air,  while  the  air  is  ulti- 


METHODS  OF   VENTILATION.  427 

mately  sent  out  into  the  open  at  a  speed  so  low  that  practically  no 
resistance  is  experienced. 

The  Shutter  Regulator. — The  volumes  which  these  fans  will 
produce  vary  directly  as  the  speed  of  their  rotation,  and  their 
manometric  depression  varies  as  the  square  of  the  speed  of  rota- 
tion. Though  they  may  be  run  at  any  speed  at  will,  the  efficiency 
of  the  fan  materially  decreases  when  the  speed  of  the  tips  of  the 
blades  exceeds,  to  a  great  degree,  5000  feet  per  minute,  or  is  less 
than  this  quantity.  The  rate,  however,  which  is  regarded  as 
normal  is  4000  feet  of  peripheral  velocity  per  minute.  Below  or 
above  the  normal  speed  a  loss  of  velocity  ensues  in  the  discharg- 
ing air,  which  alternately  is  expelled  into  the  chimney  or  carried 
with  the  blades  into  the  fan,  there  to  repeat  its  circuit.  The 
discharge  is  frequently  followed  by  a  vibration  in  the  fan,  to 
remedy  which  the  sliding  shutter  (ab,  Fig.  159)  is  introduced. 
Its  correct  position  is  only  known  by  experiment  in  each  indi- 
vidual case,  determined  by  the  point  at  which  the  throbbing 
ceases  with  the  given  speed.  Though  now  universally  placed 
in  all  fans  the  " anti- vibration  shutter"  originated  with  the 
Walker  fan. 

Influence  of  Shape,  Dimension,  and  Speed  of  Fan  upon  its 
Capacity. — The  ratio  between  the  speed  and  the  yield  seems  to  be 
as  the  4th  power  of  the  speed  to  the  5th  power  of  the  yield.  The 
relation  between  the  volume  resistance  of  the  fan  and  the  power 
required  to  drive  it  depends  upon  the  resistance  of  the  mine 
which  it  supplies.  These  ratios  have  not  been  accurately  for- 
mulated. 

Numerous  experiments  have  been  conducted  upon  centrifugal 
ventilators  with  the  view  to  determining  the  influence  which  the 
various  dimensions  of  the  fan  and  shapes  of  its  parts  will  have 
upon  its  performance;  and  the  following  conclusions  are  cited 
from  the  results  of  the  tests  made  by  R.  Van  A.  Norris,  Wilkes- 
barre,  Pa.,  upon  25  fans,  as  the  influence  of:  "ist.  The  diameter 
on  their  performance  seems  nil;  the  only  advantage  of  large 
fans  being  in  greater  width  and  a  lower  speed  required  of  the 
engines.  2d.  Width  upon  efficiency  is,  as  a  rule,  small.  3d.  Shape 


428  MANUAL  OF  MINING. 

of  blades  shows  that  the  back  curvature  is  better,  and  dimin- 
ishes the  vibration.  4th.  Shape  of  casing  is  considerable. 
The  proper  shape  would  be  one  of  such  form  that  the  air  be- 
tween each  pair  of  blades  would  constantly  and  freely  discharge 
into  the  space  between  the  fan  and  casing,  the  whole  being 
swept  to  the  &vas£e  chimney.  A  large  spiral,  beginning  at  or  near 
the  point  of  cut-off,  gives  in  every  case  a  large  efficiency.  5th. 
The  shutter  on  the  fan  is  beneficial.  The  exit  area  can  be  reg- 
ulated to  suit  the  varying  quantity  of  air,  and  prevent  re-entries. 
6th.  Speed  at  which  the  fan  is  run.  The  efficiency  is  high  if  the 
peripheral  velocity  is  large." 

Automatic  Speed  and  Pressure  Recorders.  —  Some  States 
require  self -recorders  on  all  fans,  by  which  the  number  of  revolu- 
tions of  the  fan  shall  be  registered  every  hour  and  such  data  to 
be  taken  and  reported.  In  other  States  also  is  required  an  auto- 
matic regulator  for  the  water-gauge.  The  speed- registers  are 
generally  constructed  of  a  metal  pedestal  erected  on  blocks  at 
the  side  of  the  fan  or  engine-shaft,  a  small  vertical  shaft  to  which 
a  governor  is  attached.  A  small  cog-wheel  on  the  lower  end 
geared  to  a  large  driver  on  the  fan  or  engine-shaft  communicates 
the  speed  to  the  governor,  which,  by  a  system  of  leverage,  raises 
or  lowers  the  arm  to  which  is  attached  a  pen  that  presses  against  a 
paper  dial  held  in  position  by  a  light  case  of  sheet  brass.  The 
higher  the  speed  of  the  fan,  the  more  will  the  governor  raise  the 
lever,  and  consequently  the  pen  register.  The  time  is  recorded 
by  a  clock  to  whose  shaft  the  dial  case  is  attached.  In  other 
devices  the  dial  case  is  a  cylinder  in  which  is  rolled  a  sheet  of 
paper  turning  on  a  horizontal  axis,  which  is  also  the  continuation 
of  the  shaft  of  the  clock.  These  instruments  perform  the  work 
expected  of  them  with  great  satisfaction. 

Comparison  of  Fan  and  Furnace. — The  ventilation  current 
from  a  fan  is  affected  by  a  low  barometer  or  a  high  temperature. 
Either  one  requires  an  increased  degree  of  rarefaction  just  as 
from  a  furnace.  Moreover,  as  the  depth  of  the  mine  increases, 
the  work  devolving  upon  the  fan  proportionally  increases,  be- 
cause normally  the  air  becomes  denser;  with  every  additional 


METHODS   OF   VENTILATION.  429 

thousand  feet  of  depth,  an  increased  rarefaction  or  depression  of 
0.4  inch  of  water-gauge  is  necessary.  Compared  with  the  fur- 
nace its  efficiency  decreases  with  the  depth  of  the  upcast  until, 
at  a  certain  depth,  it  becomes  an  open  question  'between  the 
relative  merits  and  demerits  of  fan  and  furnace,  as  to  which  will 
be  the  more  economical.  For  shallow  works  the  exhaust-fan 
undoubtedly  takes  precedence.  At  the  depth  of  a  thousand  feet 
a  large  furnace  will  equal  a  very  imperfect  fan,  consuming  20 
Ibs.  of  fuel  per  hourly  horse-power;  a  good  fan  and  condensing 
engine  will  be  cheaper  than  a  furnace  down  to  the  depth  of  about 
4000  feet.  Taking  cognizance  of  the  objections  to  the  furnace, 
it  must  also  be  borne  in  mind  that  machine  ventilators  are  sub- 
ject to  serious  objections,  since  during  the  time  of  their  repair 
the  mine  must  remain  un ventilated,  whereas  with  a  furnace,  after 
its  fire  has  been  extinguished,  a  considerable  circulation  will  still 
continue  in  the  upcast  for  some  time.  Auxiliary  ventilating 
appliances  should  be  supplied  against  any  emergency  which 
arises  during  the  repair  of  the  fan. 

The  Theory  of  the  Fan. — The  theoretical  depression  of  a  fan, 
the  difference  between  the  pressure  of  the  entering  air  and  the 
discharged  volume,  is  represented  in  terms  of  a  motive  column 
of  the  density  of  the  inflowing  air.  An  ideal  ventilator  will 
produce  a  depression  which  is  twice  the  height  created  by  the 
tangential  speed  of  the  tips  of  the  blades.  If,  then,  H  be  the 
height  of  the  motive  column  due  to  the  velocity,  T,  of  the  fan- 

J-2 

tips  in  feet  per  second,  H  will  equal  — ;   but  imperfections  of 

detail  prevent  such  an  initial  depression  being  attained,  and 
representing  them  by  a  coefficient  K,  which  is  always  less  than 
unity, — reaching  0.85  in  Guibal  fans,  but  more  often  being 
below  0.6  in  the  average  construction  of  fans, — the  fan  approaches 
an  ideally  perfect  one  when  K  approximates  to  unity.  The 
yield  of  the  fan  then  in  barometric  depression,  or  its  useful  effect, 

KT2 
is   H  = .     Various   enfeebling   causes   reduce    the    capacity 

o 

of  the  fan  to  'determine  the  value  for  K.    The  quantity  of  air 


4,)0  MANUAL   OF  MINING. 

which  passes  through  an  orifice  is  equal  to  the  product  of  the 
area  and  the  velocity  when  no  friction  is  encountered;  but  when 
any  fluid  flows  through  an  orifice  in  a  plane-  surface  a  considerable 
diminution  of  the  discharge  takes  place,  because  of  the  contraction 
of  the  steam  resulting  from  the  convergent  flow.  The  coefficient 
corresponding  to  this  contraction,  known  as  the  vena  contracta, 
is  0.65;  hence  with  a  given  velocity,  T,  and  a  head,  H,  under 
the  conditions  modified  by  the  coefficient  K  as  above,  the  dis- 
charge of  air  per  second  will  become 


Hence  it  is  evident  that  if  the  capacity  of  the  mine  is  such 
that  it  is  incapable  of  delivering  to  the  fan  the  volume  of  air 
equal  to  the  body  capacity  of  the  latter  at  a  given  speed,  the 
frictional  resistances  encountered  in  the  mine  will  reduce  the 
efficiency  of  the  fan  by  some  quantity  which  is  usually  com- 
prehended in  the  symbol  a,  representing  the  area  of  the  mine's 
"equivalent  orifice"  in  square  feet  Experiments  have  demon- 
strated that  when  a  is  20  square  feet,  only  65,000  cubic  feet  of 
air  are  obtained  per  minute  for  the  fan  peripheral  speed  of  5000 
feet  per  minute  ;  but  when  the  mine  resistances  have  been  reduced 
until  its  "equivalent  orifice"  is  as  large  as  100  square  feet,  280,000 
cubic  feet  of  air  are  obtained  from  the  same  speed  of  fan. 
Having  the  value  for  this  friction,  which  in  earlier  days  was 
known  as  the  temperament  of  the  mine,  we  arc  enabled  to 
understand  the  conditions  under  which  the  ventilator  is  working 
and  to  provide  for  them. 

The  Equivalent  Orifice  of  the  Fan,  which  is  designated  by 
the  symbol  o,  may  be  determined  in  a  like  manner.  It  measures 
or  represents  the  orifice  in  a  thin  plate  which  offers  such  a  resist- 
ance to  the  flow  of  the  current,  Q,  as  is  equal  in  effect  to  the 
aggregate  resistances  encountered  within  the  fan  from  its  imper- 
fections. If  H  is  the  theoretical  depression  which  the  fan  should 
produce  when  moving  at  a  tangential  speed,  T,  per  second,  and 
h  represents  the  actual  or  the  effective  depression  which  is  pro- 


ELEVATION 
FIG.  1 60. — Details  of  a  Fan  House. 


43* 


43  2  MANUAL  OF  MINING. 

duced  upon  the  air  as  measured  by  the  water-gauge,  then  H—  h 
is  the  head  wasted  by  the  fan  in  its  construction  and  may  be 
represented  by  &0,  which  measures  the  head  corresponding  to 
the  equivalent  orifice.  In  large  fans  its  value  varies  from  16 
to  80  square  feet. 

The  head  lost  in  the  fan,  represented  by  h0,  is  equal  to  H—  ht 
the  velocity  due  to  which  may  be  determined  by  the  expression 


Aslhe  value  of  h0  approaches  zero  and  that  of  h  approaches 
H,  the  fan  approaches  an  ideally  perfect  ventilator.  The  actual 
velocity  through  the  orifice  of  entry  is  0.657;,,,  whence  the  area 
of  the  orifice  o,  which  equals  the  quantity  flowing  per  second, 
divided  by  the  velocity  of  the  flow,  has  the  following  value: 


The  density  of  water  being  833  times  that  of  air,  the  ratio 
between  the  water-gauge  reading  and  the  height  of  the  motive 
column,  H,  is  i  :  833.  To  convert  the  water-gauge  reading 
to  a  height  H  of  air-column  of  equal  weight  in  feet,  the  height 
of  the  water-gauge,  m,  in  inches  is  multiplied  by  69.4. 

The  ratio  between  the  lost  head  in  the  ventilator  and  the 
effective  head  represented  by  the  water-gauge  is  expressed  in 
the  equation 


whence 


The  Relation  between  Mine  and  Fan.  —  The  quantity  of 
flow  through  a  fan,  depending  on  the  relations  which  the  area 
of  the  mine  airways  and  the  condition  of  their  rubbing  surface 
bear  to  the  mechanical  condition  of  the  fan,  is  manifestly  depend- 


METHODS  OF   VENTILATION. 


433 


SECTION  ON  LINE  A-B 

PLAN 
FlG.  161. — Plan  of  the  Fan  House  and  the  Curve  of  the  Fan  Housing. 


434  MANUAL  OF  MINING. 

ent  upon  a  proper  ratio  of  a  to  o,  which  ratio  may  be  expressed 
as  the  "appropriateness  of  the  fan  to  the  mine."  When  this 
value  is  equal  to  or  greater  than  unity,  the  fan  would  be  too 
small  for  the  mine,  and  it  is  questionable  whether  any  air  would 
flow  under  those  conditions.  As  the  ratio  becomes  smaller,  the 
conditions  become  more  favorable  for  the  fan.  When  approxi- 
mating a  ratio  of  0.3  the  orifice  of  discharge  of  the  fan  is  to  be 
considered  as  having  a  fair  working  ratio.  More  air  is  obtained 
by  a  given  fan  and  at  a  given  velocity  when  a  is  large  than  when 
•a  is  small,  for,  no  matter  how  well  constructed  the  fan  may  be, 
it  cannot  provide  a  quantity  equal  to  its  body  capacity  unless 
the  mine  can  pass  this  quantity.  The  effective  work  done  upon 
the  air  is  less  in  the  latter  case  than  in  the  former  for  a  given 
volume  of  air.  The  mechanical  work  of  centrifugal  force  is 
o.oooo34o(r2— F22).  In  this  T  is  the  circumferential  velocity 
and  V2  is  the  absolute  velocity  at  expulsion,  due  to  compression 
from  centrifugal  force.  As  F2  increases,  so  the  work  on  the 
departing  air,  and  proportionately  the  effective  work,  decreases. 
The  use  of  the  funnel-chamber  reduces  this  quantity  to  £  or  -5,  and 
the  work  lost  to  4  or  5  per  cent. 

The  Efficiency  of  the  Fan. — The  manometric  efficiency  is 
the  ratio  between  the  effective  pressure,  P,  and  that  due  to  the 
centrifugal  velocity,  while  the  mechanical  efficiency  is  that  of  the 
Horse-power  in  the  air  to  the  engine-duty.  With  fan  properly 
constructed,  the  efficiency  approximates  about  6  per  cent.  In 
experimentally  measuring  the  efficiency  of  a  fan,  it  is  customary 
to  determine  the  dynamometric  resistance  and  internal  friction 
when  its  orifices  have  been  cut  off  from  any  communication 
with  the  mine,  the  air  being  then  drawn  from  the  atmosphere  and, 
after  passing  through  a  fan,  expelled  at  its  throat.  Counting 
the  rate  of  revolution  and  estimating  the  volume  of  air  which 
is  moved,  the  power  necessary  to  overcome  this  friction  is  deter- 
mined and  expressed  in  feet  of  the  air-column  whose  weight 
equals  the  aggregate  friction.  This  quantity,  divided  by  the 
theoretical  head  corresponding  to  the  velocity  of  the  fan,  deter- . 
mines  its  efficiency  under  the  conditions  named.  The  fan  is 


METHODS  OF   VENTILATION.  435 

giving  its  maximum  efficiency  when  "its  body  capacity  just 
exceeds  the  quantity  the  mine  will  pass  at  a  gauge  pressure,  F, 
due  to  the  speed  of  rotation  of  the  fan." 

The  Design  of  a  Fan. — In  designing  a  fan  for  a  given  mine, 
the  essential  elements  are  Q  and  m.  These  given,  the  diameter, 
the  peripheral  speed,  and  the  length  and  width  of  blades,  as 
well  as  the  direction  of  their  inclination,  must  be  determined 
by  the  engineer.  As  to  diameter,  it  may  be  said  that  the  slow- 
running  fans  are  regarded  as  cumbersome  and  costly,  requiring 
expensive  foundations.  Large  fans  may  be  run  at  a  lower  rate 
of  revolution  and  produce  the  same  tangential  speed  as  would  a 
fan  of  small  diameter.  Insomuch  as  speed  is  the  important 
factor  in  the  construction  of  ventilators,  due  consideration  must 
be  given  to  this  question,  which  is  determined  by  local  condi- 
tions of  place,  economy,  and  mechanical  simplicity.  A  con- 
venient rate  of  revolution  for  a  fan  directly  connected  with  the 
engine  is  about  60  per  minute.  The  body  capacity  of  the  fan 
should  be  large  enough  to  maintain  the  required  pressure,  P, 
without  great  variations  in  speed.  Though  the  practice  of 
European  engineers  tends  towards  the  rate  of  tangential  speed 
which  represents  5000  feet  per  minute  and  over,  in  this  country 
4000  may  be  considered  as  normal.  In  any  event,  if  the  calcu- 
lation and  design  be  made  on  the  assumption  of  either  normal 
speed,  it  will  be  possible,  when  an  emergency  arises,  to  increase 
the  speed  sufficiently  to  give  a  volume  nearly  one  tenth  greater 
than  the  normal  quantity.  Moreover,  when  the  mine  is  dry 
and  dusty,  it  will  be  possible  to  turn  the  whole  volume  of  the 
excess  into  any  or  each  single  split,  through '  which  it  may  be 
drawn,  clearing  away  fine  dust  and  moisture. 

The  entry  for  the  air  should  be  made  easy  and  large,  prefer- 
ably divided  into  two  inlets,  one  on  each  side,  with  a  diaphragm 
to  prevent  the  currents  from  conflicting.  This  necessitates  a 
wide  fan,  which,  however,  gives  a  volume  proportionately  greater 
than  what  is  to  be  had  from  a  single  fan  with  a  single  large  inlet. 

The  length  of  the  blades  of  the  fan  should  be  only  a  little 
greater  radially  than  the  difference  between  the  radii  of  the 


43<5  MANUAL  OF  MINING. 

fan  and  its  Inlet.  With  a  large  inlet  the  blade  necessarily  is 
shortened,  and  when  pressure  is  desired  the  blade  length  should 
be  increased  to  as  large  a  quantity  as  possible  by  providing  two 
inlets.  Notwithstanding  that  the  width  of  the  fans  is  much  greater 
than  would  be  obtained  by  substitution  in  the  formulae  follow- 
ing, it  is  certain  that  the  latter  dimensions  correspond  to  a  greater 
efficiency.  M.  G.  Hanarte  concludes  that  "the  Guibal  fan  has 
always  been  eight  or  nine  times  too  wide,  and  the  Capell  is  nearly 
as  bad." 

The  shape  of  the  blade  should  be  such  as  would  present  to- 
the  circumference  of  the  outlet  an  inclination  following  the 
resultant  of  the  movement  of  rotation  and  of  the  movement  of 
the  air  penetrating  the  spaces  between  them.  The  blades  of 
open-running  fans  curve  backward.  The  backward  curvature 
is  conceded  to  give  a  freer  delivery,  and  the  forward  curvature 
at  the  tips  a  higher  water-gauge  pressure.  The  number  of  blades 
is  seemingly  a  matter  of  indifference,  though  the  limit  may  doubt- 
less be  determined  by  the  inevitable  friction  produced  by  the 
excessive  surface  of  contact  when  too  numerous.  The  friction 
varies  as  the  cube  of  the  section  of  space  between  two  vanes. 
As  to  the  shape  to  be  given  to  the  casing,  it  will  be  noticed  that 
the  original  Guibal  fan  had  no  spiral,  the  tips  of  the  blades  re- 
volving but  two  inches  clear  of  the  casing,  and  the  spiral  enlarge- 
ment beginning  at  the  angle  of  about  67°  30'  from  the  lower 
vertical  radius.  Those  fans  presenting  a  large  spiral  beginning 
at  or  near  the  cut-off  and  increasing  about  six  inches  for  each, 
45°  up  to  275°,  and  thence  widening  by  an  increasing  increment 
to  the  evasee  chimney,  appear  to  give  larger  efficiency  by  allowing 
for  the  slackening  of  the  speed  of  the  air,  and  discharge  the  air 
with  less  energy  at  the  exit.  M.  G.  Hanarte  concludes  that  the 
spiral  envelope  is  not  necessary. 

Below  are  given  formulae  for  the  computation  of  dimensions 
of  a  Guibal  fan  in  accordance  with  the  data  indicated  above. 
All  dimensions  are  in  feet. 

.D=diameter  of  the  fan  between  the  blade-tips:    Q^-2oo=Z>2; 
/==  length  of  the  blades  in  feet; 


METHODS  OF   VENTILATION. 

r=their  radial  length  =2. 6im; 

x  =  their  width  =A  .-*-  27:5; 

A  =  aggregate  area,  of  the  one  or  two  inlet-ports  in  square  feet 
(radius  of  each  central  inlet,  s)  =Q-r- 1300; 

JV=number  of  revolutions  per  minute; 

r=peripheral  speed  of  fan  per  second  =1^-^19.0985; 

F=  theoretical  velocity  per  second  due  to  head  H; 

vl  =  velocity  of  the  centre  of  gyration  of  air-column  between* 
the  blades  =7t(D-r}N+  60; 

p=  radius  of  gyration  of  the  mass  of  air=^(Z)— r). 

W=  weight  of  the  unit  of  revolving  air-column  per  foot  of  fan- 
width  =o.o766r;  \ 

F=  centrifugal  force  of  the  air-mass  in  Ibs.  per  square  foot  of  dis- 
charge area  or  of  the  housing  =TF'i>,2-^7rg=o.ooo7578rz;12;. 

F2=velocity  of  air  discharge  per  second  =>J- ™'°°°  r 

Z=  minimum  area  of  discharge-port  =A  -r-  2 ; 

h0=fa.n  resistance,  measured  in  feet  of  head,  =H—h\ 

O=area  of  orifice  offering  a  resistance  to  the  flow  of  Q  cubic 

feet  of  air  per  minute,  equal  to  that  of  the  fan; 
q=  quantity  of  air  discharged  by  the  fan  per  second  in  cubic 

feet=o.6sF2Z; 
Q  —  quantity  of  air  discharged  by  the  fan  per  minute  in  cubjc 

feet  =60?; 

w=mine  resistance  in  inches  of  water-gauge; 
P=mine  resistance  in  pounds  per  square  foot  =5. 184  m. 

In  Figs.  160  to  162  are  illustrated  the  details  of  the  ordinary- 
pattern  of  fan  which  is  designed  in  accordance  with  the  con- 
ditions indicated  above.  As,  fortunately,  neither  the  Guibal 
fan  nor  the  shutter  is  subject  to  patent,  the  working  drawings 
here  given  may  aid  the  construction  engineer. 

When  the  conditions  are  satisfied  by  the  revolution  of  the 
fan  of  proper  proportions,  the  centrifugal  pressure  of  the  fan- 
should  produce  a  depression,  F,  equal  to  or  exceeding  P,  the 
mine  resistance,  in  order  that  the  requisite  discharge  through 


438 


MANUAL  OF  MINING. 


Flo.  162. — Details  of  the  Fan. 


METHODS  OF   VENTILATION, 


43? 


the  outlet  should  equal  the  desired  quantity  Q.  When  it  is 
discovered  that  the  volume  of  discharge  is  deficient,  the  fan 
dimensions  D  or  r  should  be  enlarged  or  the  rate  of  revolution 
increased.  Below  is  a  brief  table  indicating  the  theoretical 
water-gauge  depression  in  inches  for  the  corresponding  periph- 
eral speeds  in  feet  per  second. 


Speed  in 
Feet 
of  the 
Periphery 
per  Second. 

Depression 
in  Feet 
of  Air- 
column. 

Depression 
in  Inches 
of  Water- 
gauge. 

Speed  in 
Feet 
of  the 
Periphery 
per  Second. 

Depression 
in  Feet 
of  Air- 
column. 

Depression 
in  Inches 
of  Water- 
gauge. 

T 

U 

m 

T       'r> 

M  ' 

m 

30 

2  7  -,95 

0.411 

75 

I74-.69 

2.569 

35 

38.04 

Q-559 

80 

I98-75 

2.922 

40 

49.69 

0.731    • 

85 

224.38 

3-299 

45 

62.82 

0.924 

9P     i 

25I-55 

3-699 

5° 

77-63 

i,.  141 

95       , 

280.28 

4.122 

55 

93-94 

1.381 

100 

310-55 

4-567 

60 

in.  80 

i  .  644 

i°5 

342.39 

5-035 

65 

131.18 

i  .929 

IIO 

375-77     , 

5-526 

70 

152.17 

2.238 

115 

410.71 

6.039 

EXAMPLE. — Required  the  dimensions  of  a  fan  to  provide  125,000  cubic 
feet  of  air  against  a  mine  resistance  of -2.5  inches. 

At  a  normal  rate  of  65  feet  per  second,  the  diameter  becomes  25  feet;  the 
area  of  the  inlets  is  96  square  feet,  the  diameter  being  n  feet;  the  radial 
length  of  the  blade  is  6.5  feet;  the  minimum  \width  of  the  blade  is  to  be  2.8 
feet. 

As  £=9.25  feet  and  ^  =  51  feet,  the  centrifugal  pressure,  F,  becomes  12.7 
pounds  per  square  foot  of  radial  column;  and  the  velocity  of  discharge  31.6 
feet  per  second,  which  with  a  minimum  area  of  discharge-port,  Z,  of  48  square 
feet,  would  furnish  less  than  60,000  cubic  feet.  The' mine  resistance  exceeds 
the  standard  allowance  of  i  inch  of  water  gauge  for  each  100,000  cubic  feet 
of  air.  The  mine  airways  should  be  enlarged  or  the  fan. operated  at  a  higher 
speed.  An  increased  rate  of  70  revolutions  per  minute  will  produce  a  ventilat- 
ing pressure  of  22  pounds  "per 'square  foot.  The 'blades -may  be  lengthened 
and  two  inlet  orifices  be  provided,  each  of  48  square  feet  in  area. 

At  the  peripheral  velocity,  T,  of  91.66  feet  per  second  the  theoretical  head 
of  discharge  is  261  feet.  But  the  effective  head,  h,  against  which  the  fan  is 
operating,  measured  by  the  water-gauge,  is  166.66  feet.  Under  the  condi- 
tions of  operation,  therefore,  the  loss  of  head,  h0,  in  the  fan  is  94.34  feet;  since 
the  equivalent  orifice  of  the  mine  is  21.2  square  feet,  the  equivalent  orifice  of 
the  fan,  O,  is  41.2  feet.  The  ratio  of  a  to  O,  n»rly  one  half,  represents  a  fair 
working  ratio  of  appropriateness  of  fan  to  mine. 


44°  MAXUAL  OF  MINING. 

Electrically  driven  blowers  and  ventilating-fans  may  be  con- 
nected either  by  belt,  gearing,  or  mechanical  coupling.  The 
compound-wound  D.  C.  machine  is  preferred  for  fans  on  account 
•of  its  good  speed  regulation  and  freedom  from  racing. 

Alternating-current  induction-motors  of  the  squirrel-cage  type 
would  give  equal  satisfaction. 

REFERENCES. 

Curvature  of  Fan  Blade  and  Design,  M.  &  M.,  Vol.  XX,  165;  Fan  Effi- 
ciency, M.  &  M.,  Vol.  XXIII,  406;  Location  of  Fan,  M.  &  M.,  Vol.  XXIII, 
22;  Fan  Ventilation,  M.  &  M.,  Vol.  XXIII,  260-71;  Laws  Governing  Useful 
Work  of  a  Fan,  H.  Heenan  and  William  Gilbert,  Coll.  Guard.,  April  1897, 
720;  and  M.  G.  Hanarte,  Coll.  Guard.,  Mar.  1898,  552;  Descriptive  Lecture 
on  Fans,  C.  M.  Percy,  Coll.  Mgr.,  1894,  56;  Fan  Volume  Recorder,  M. 
Murge,  Coll.  Guard.,  April  12,  1900;  Modern  Fans,  Charles  H.  Innes,  Eng., 
Nov.  14,  1902;  and  Theory  and  Practice  of  Fan  Design,  Heating  and  Ven- 
tilation, Serial,  1903  and  1904. 

Underground  Temperatures  in  Relation  to  Deep  Mining,  F.  C.  Meachem, 
Coll.  Guard.,  April  7,  1903;  The  Practical  Scope  of  the  Equivalent  Orifice 
Theory,  H.  W.  Halbaum,  Coll.  Guard.,  Aug.  17,  1900. 

Improvised  Ventilating  Machinery,  C.  M.  Neyrick,  Min.  Sci.  Press,  April 
'  n,  1903;  Ventilation  Furnace,  Illustrated,  Pa.  Mine  Insp.,  1891,  404,  190. 

Root  Blower,  Haswell,  449;  To  Compute  Elements  of  Fan,  Haswell,  448; 
Fan  Temperament,  Clifford,  Am.  Mfr.,  Feb.  19,  1897,  264;  Formulae,  Coll. 
Guard.,  Mar.  12,  1897,  489;  Fan  Work,  Coll.  Guard.,  Mar.  19,552;  Mano- 
metric  Efficiency,  Fed.  Inst.  M.  E.,  111.;  Fan  Tests  by  Van  A.  Norris,  A.  I. 
M.  E.,  XX,  637;  The  Comparative  Consumption  of  Fuel  by  Ventilating  Fur- 
naces and  by  Ventilating  Machines,  J.  J.  Atkinson,  N.  E.  I.,  VI,  135. 

Guibal  Fan  Compared  with  Dynamo,  Coll.  Guard.,  Vol.  LXXXIV,  847; 
Comparative  Tests  of  Fans,  Coll.  Guard.,  LXXIX,  218;  Double  Ventilation 
Fans,  Coll.  Guard.,  LXXXI,  127. 

Equivalent  Orifice  of  Mine,  Table  for  Finding,  Coll.  Guard.,  Jan.  1897, 
.168;  Speed  Indicators,  Pa.  Mine  Insp.,  1886,  22,  24;   1891,  89. 
'  '•'  JA 

ih  to 
•' 

:>ifr    . 

• 


: 

• 
• 

' 

CHAPTER  XIV. 

DISTRIBUTION  OF  THE  AIR. 

The  Effective  Motive  Column. — It  has  been  assumed  that 
the  work  done  upon  the  air  is  totally  effective  in  the  mine;  that, 
with  a  given  M  and  P,  the  calculated  quantity  of  air  is  obtained 
without  loss;  that  the  momentum,  once  imparted  to  the  air, 
would  be  available  to  carry  it  through  the  mine  and  out.  This 
is  not  so.  Friction  consumes  some  of  the  kinetic  energy;  only 
a  fractional  part  of  the  theoretical  velocity  is  realized  in  practice. 
The  rough  sides  of  the  galleries  and  rooms,  their  sharp  corners, 
and  the  diminished  areas  offer  resistances  to  the  passage  of  the 
current  that  consume  often  90  per  cent  of  the  power.  More- 
over, the  subtle  air  under  pressure  seeks  to  escape  at  every  oppor- 
tunity, and  some  portion  of  the  precious  fluid  is  lost  into  the  goaf, 
through  doors  and  at  crossings.  A  certain  mine  theoretically 
required  a  pressure  of  but  1.2  Ibs.  per  foot  to  give  rise  to  its  cur- 
rent, yet  the  friction  was  such  that  u.8  Ibs.  were  actually  neces- 
sary to  create  the  velocity.  Not  infrequently  the  ratio  between 
M  (to  which  the  generation  of  the  final  velocity  at  the  top  is 
due)  and  M',  the  head  actually  necessary  to  overcome  resist- 
ances, is  as  low  as  i :  18.  In  other  words,  only  5.5  per  cent 
of  the  work  done  upon  the  air  is  usefully  expended.  Any  means 
of  reducing  this  loss  is  to  be  welcomed. 

The  Frictional  Resistance  of  Air  in  Mines. — Let  us  examine 
into  the  laws  governing  the  movement  of  fluids,  that  we  may 
reduce  this  friction  to  a  minimum,  and  obtain  salubrity,  safety, 
and  economy  with  the  least  outlay.  The  air  which  enters  the 
mine  from  the  downcast  is  distributed  to  the  rooms  and  chambers 
hi  proportions  varying  with  their  several  needs;  or  the  current 

441 


442  MANUAL  OF  MINING. 

as  one  mass  sweeps  through  the  main  way,  along  working  faces, 
thence  by  return  airway  over  the  furnace  or  to  the  fan.  The 
resistances  encountered  depend  upon  the  ratio  of  the  area  of 
the  surface  rubbed  to  the  area  of  the  conduit,  and  upon  the  coef- 
ficient of  air  friction  against  rock.  A  satisfactory  value  for  the 
coefficient  has  not  been  obtained:  the  records  of  experiments 
show  it  to  vary,  as  in  water,  according  to  the  nature  of  the  conduit 
and  the  velocity  of  the  flow.  The  coefficient  varies  with  the 
nature  of  the  rubbing-surface,  and  consequently  differs  in  various 
air-  passages  of  the  same  mine;  nevertheless,  the  numerous  ex- 
perimenters have  announced  values  for  the  coefficient  of  friction 
of  air  in  mines  for  each  foot  of  rubbing-surface  and  for  a  velocity 
of  one  foot  per  minute  as  varying  between  0.000000008585 
and  0.0000000219,  with  an  accepted  average  of  o.oooooooi  Ib. 
This  value  for  /,  the  coefficient,  is  measured  in  the  pounds  pres- 
sure per  square  foot.  Measuring  the  height  of  an  air-column 
in  decimals  of  a  foot,  the  value  for  the  coefficient  of  friction 
/'  has  been  found  to  be  0.00000010635  to  0.00000026881,  the 
two  extremes  of  values  for  /'  being  heads  of  air-column  corre- 
sponding to  the  values  given  for  dynamic  pressures,  /.  In  other 
words,  an  area  of  i  square  foot  of  rubbing-  surf  ace  presents  to 
the  passage  of  an  air-current  moving  at  i  foot  per  minute  a 
frictional  resistance  of  o.oooooooi  Ib.,  or  an  equivalent  resistance 
of  0.000000125  feet  of  head  of  air-column. 

Let  /  be  the  length,  m  the  perimeter,  and  a  the  area  of  the 
gangway,  through  which  the  air  is  coursing  at  v  feet  per  minute, 
and  the  rubbing  friction  is  found  experimentally  to  be  jlmv2. 
Imagine  a  piston,  fitting  air-tight  in  the  passage;  to  just  move 
it  against  the  resistance  requires  the  expenditure  of  a  force  pa, 
in  units  of  pounds  and  square  feet.  Therefore  the  loss  of  power 

due  to  friction  is  pa  =  jlmv*,  and  the  loss  of  head  in  feet,  p=  —  —  , 


or  in  pounds, 


j'lmv* 


Shape    and    Dimensions   of  Airways.  —  Cognizance   must   be 
taken  of  these  frictional  resistances  by  ascertaining,  for  various 


DISTRIBUTION   OF    THE  AIR.  443 

portions  of  the  mine,  the  elements  involved  and  comparing  the 
value  for  p,  thus  ascertained  for  .a  given  mine,  with  that  usually 
considered  as  the  standard  of  comparisons.  The  normal  mine 
should  be  capable  of  delivering  100,000  cu.  ft.  of  air  to  the  men 
throughout  its  working  without  causing  a  frictional  resistance 
greater  than  5.2  Ibs.  per  square  foot  in  the  aggregate.  Should  it 
be  ascertained  that  the  mine  shows  a  larger  water-gauge,  it  is 
evident  that  there  is  room  for  improvement. 

The  lines  along  which  improvement  may  be  had  involve  an 
increase  in  the  area  of  the  .cross-section,  or  a  change  of  shape, 
such  as  to  give  a  greater  flow  with  less  frictional  rubbing- surf  ace. 
The  conditions  under  which  the  airways  are  driven  prevent  any 
reduction  in  the  coefficient  of  friction.  The  friction  increases 
with  the  square  of  the  velocity,  and  in  order  to  increase  the  air- 
current  the  velocity  may  be  increased,  but  this  increases  the 
frictional  resistances  rapidly.  Assuming  the  value  for  /  of 
o.ooooooi,  then  the  quantity  that  will  flow  through  i  sq.  ft. 
of  area  of  passage  at  different  velocities,  as  stated  below,  in  a 
length  of  one  mile,  will  be  as  indicated,  and  with  corresponding 
resistance,  p,  as  follows: 

Feet.  Cubic  Feet.  Lbs.  per  Square  Foot. 

Velocity    i,  quantity  Q=  60;  then^  =  0.76 

"    '     2,  "       <2=i2o;  "    p=  3.06 

5,  «•       <2=3oo;  "    f =19.00 

"       10,  "       Q  =600;  "    .£  =  76.03 

The  frictional  loss  is  directly  proportional  to  the  perimeter, 
and  inversely  as  the  area.  Thus,  a  local  contraction  of  the 
air-passage  by  the  use  of  the  partly  open  door,  or  the  accidental 
accumulation  of  a  pile  of  waste,  or  gangue,  will  materially  dimin- 
ish the  volume  of  air  passing  through  it  and  increase  in  a  large 
measure  the  resistance  to  such  flow  as  does  occur.  It  is  there- 
fore desirable  to  select  such  a  shape  for  the  airway  as  will  make 
it  as  spacious  as  circumstances  will  permit  consistent  with  the 
dimension  of  the  rubbing- surf  ace.  The  circular  form  most  nearly 
meets  these  requirements,  but  we  are  restricted,  as  a  rule,  to  the 


444 


MANUAL  OF  MINING. 


rectangular  or  the  trapezoidal  cross-section.  Aside  from  the 
theoretical  consideration,  the  form  of  airway  is  modified  by  prac- 
tical conditions.  In  American  mines  the  circular  airway  is 
rarely  adopted,  and  only  in  very  rare  instances  is  the  Stanley 
header  used  as  in  England.  There  a  circular  lining  follows 
the  machine  very  closely,  and  the  proper  form  can  be  retained  in 
the  finished  airway.  The  expense  of  making  this  form  pre- 
Tents  its  common  use  in  America.  The  more  usual  form  is 
one  having  a  semicircular  arch  roof  and  smooth  walls  as  being 
the  nearest  approach  to  the  ideal  cross-section. 

As  showing  the  comparisons  between  the  resistance  afforded 
by  a  square,  trapezoidal  road  and  the  round  roadway  of  various 
dimensions  and  velocities,  as  stated  below,  the  following  table 
will  be  of  service.  It  shows  the  pressure  in  pounds  per  square 
foot  due  to  the  flow  of  air  through  one  mile  of  airway  with  the 
coefficient  of  o.ooooooi  Ib.  per  square  foot. 


At  Velocity  of 

At  Velocity  of 

At  Velocity  of 

Dimensions 
of 

Diameter 
of 

i  Foot  per 
Second. 

S  Feet  per 
Second. 

i  o  Feet  per 
Second. 

t 

Q 

P 

Q 

P 

Q 

Feet. 

Feet. 

Pounds. 

Pounds. 

Pounds, 

IX    I 

i 

.7603 

60 

19.00 

300 

76.03 

600 

2X    2 

2 

.,•3801 

240 

9-50 

1,200 

38.01 

2,400 

3X    3 

3 

•2534  ' 

540 

6-33 

2,700 

25-34 

5,400 

4X    4 

4 

.  1900 

960 

4-75 

4,800 

19.00 

9,600 

5X  5 

5 

.1520 

1500 

3.80 

7,500 

15.20 

15,000 

6X  6 

6 

.1267 

2160 

3.16 

10,850 

12.67 

21,600 

7X   7 

7 

.  1086  ' 

2940 

2.71 

14,700 

10.86 

29,400 

8X  8 

8 

•0954 

3840 

2.38 

19,200 

5-94 

38,400 

9X  9 

9 

.0844 

4860 

2.  II 

24,300 

8-44 

48,600 

10X10 

10 

.0760 

6000 

1.90 

30,000 

7.60 

60,000 

It  will  be  noted  then  that  two  galleries  having  a  cross-sec- 
tional area  of  5'Xs'  would  require  twice  as  much  power  to  carry 
the  same  amount  of  air  as  would  a  single  gallery  of  lo'Xio'. 
The  pressure  necessary  to  drive  25  cu.  ft.  of  air  per  second  through 
one  of  the  small  galleries  is  3.8  Ibs.  over  the  total  area  of  the 
road.  The  pressure  necessary  to  drive  100  cu.  ft.  per  second 
through  a  large  gallery  would  require  the  pressure  of  7.6  Ibs.  over 


DISTRIBUTION  OF  THE  AIR. 


445 


the  total  cross-sectional  area.  This  is  equal  in  amount  to  resist- 
ances of  both  small  galleries,  which,  however,  are  transmitting 
but  50  cu.  ft.  of  air  per  second. 

The  frictional  resistance  varies  with  the  length  of  the  gallery. 
The  volumes  of  air  passing  through  two  having  the  same  resist- 
ance will  be  inversely  as  the  square  root  of  their  lengths.  For 
example,  a  gallery  1600  feet  long  carrying  6000  cu.  ft.  of  air  offers 
the  same  resistance  and  consumes  the  same  amount  of  power  as 
one  of  equal  area  which  is  711  feet  long  delivering  9000  cu.  ft. 
per  minute. 

The  following  table  shows  the  length  of  roads  offering  a 
resistance  equal  to  i  inch  of  water-gauge  under  the  conditions 
given : 


Square  Airway. 

Round  Airway. 

Dimensions 
of 

Area. 

Periphery, 

Diameter 
of 

Area. 

Circumfer- 
ence, 

Length 
of  Road 
in  Yards. 

Airway. 

Airway. 

tn. 

Feet. 

Sq.  Feet 

Feet. 

•  ;  •  '•      '• 

iX   i  . 

I 

4 

I    , 

-7854 

3.1416 

481 

2X     2 

4 

8 

2 

3-I4I 

6.2832 

962 

3X   3 

9 

12 

3 

7.068 

9.4248 

1444 

4X   4 

16 

16 

4 

12.566 

12.5664. 

1926 

5X   5 

25 

20 

5 

I9-635 

15.7080 

2407 

6X   6 

36 

24 

6 

28.274 

18.8496 

2888 

7X    7 

49 

28 

7 

38.484 

21.9912 

337° 

8X  8 

64 

32 

8 

50-265 

25.1328 

3850 

9X  9 

81 

36 

9 

63.617 

28.2744 

4333 

10X10 

100 

40 

10 

78.540 

31.4160 

4814 

In  using  this  table  it  will  be  remembered  that  the  volume  of 
air  and  the  lengths  of  the  road  are  directly  proportional  to  the 
frictional  resistance,  p,  the  other  conditions  of  area  and  periph- 
ery being  the  same.  For  example,  a  gallery  which  is  1200  yards 
long,  carrying  6000  cu.  ft.,  would  offer  one  half  the  resistance 
that  a  road  2400  yards  long  would,  or,  in  other  words,  if  the  cross- 
sectional  area  be  s'Xs'i  the  water-gauge  would  read  J  inch  in 
the  short  gallery  and  i  inch  in  the  long  one.  Again,  the  road- 
way of  f  Xf,  which  is  6740  feet  long,  would  offer  a  resistance 
measured  by  a  water-gauge  of  two  thirds  of  an  inch. 


446  MANUAL  OF  MINING. 

Calculation  of  Mine  Resistance.  —  It  has  already  been 
remarked  that  the  water-gauge  measures  the  drag  of  the  air  in 
a  mine,  and  thus  serves  to  indicate  the  pressure  and  head  cor- 
responding to  the  motive  column,  M.  The  pressure  varies  from 
|  inch  for  easy  to  4  inches  for  difficult  ventilation  (from  3.9  to 
20.7  Ibs.  per  square  foot).  In  anthracite  mines  it  is  about  2 
inches.  The  motive  column,  which  is  to  just  maintain  this 
pressure  against  resistances,  should  also  be  sufficient  to  create 
a  final  or  exit  velocity  in  the  shaft.  If  the  entire  current  trav- 
erses the  mine  unbroken,  the  resistance  in  the  shaft  or  entry  is 
only  a  fractional  part  of  the  mine  friction  indicated  by  the  water- 
gauge,  and  the  following  formulae  apply  with  sufficient  accuracy: 

V^P^L  Q,J*P 

jlm'  V      jlni 

The  value  for  /  is  to  be  taken  always  in  the  same  unit  as  that 
for  p.  In  other  words,  if  the  mine  resistance,  p,  be  given  in 
pounds  per  square  foot,  the  corresponding  value  for  /  is  taken 
as  equal  to  0.0000000219;  or  if  the  value  for  p'  is  given  in  feet  of 
head  of  motive  column,  M,  the  value  for  /'  is  then  0.000000269. 

If  the  airways  in  the  mine,  the  resistances  of  which  are  to  be 
calculated  with  a  view  to  determining  the  necessary  ventilator 
pressure  to  produce  circulation,  are  all  of  the  same  dimensions, 
the  calculation  of  the  lost  pressure  may  be  made  in  one  operation 
by  proper  substitution  for  the  length,  periphery,  and  area  of 
airway  and  the  velocity  or  quantity  concerned.  The  value  of 
the  frictional  resistance,  />,  thus  engendered  in  the  mine  cor- 
responding to  the  water-gauge  height,  m,  and  of  the  velocity  of 
the  air-cuifent,  added  to  that  of  the  pressure,  p",  requisite  for 
the  generation  of  the  velocity,  determines  the  motor  pressure 
required.  Often  p"  is  very  small  compared  to  p,  and  may  be 
even  neglected  without  sensible  error;  but  when  it  is  large,  the 
actual  ventilating  pressure,  which  must  be  supplied  by  the  force- 
fan,  or  the  manometric  depression  to  be  produced  by  a  furnace, 
or  exhaust-fan,  must  be  such  as  exceeds  the  sum  of  />+  p". 

When-  the  airways  of  the  mine  are  of  various  cross-sections, 
the  resistance  offered  by  them  in  the  aggregate  must  be  deter- 


DISTRIBUTION   OF   THE  AIR.  447 

mined  by  adding  together  the  separate  values  for  p,  calculated 
for  each  differing  cross-sectional  area  and  length.  When  the 
air-current  is  "split"  into  several  smaller  branches,  and  circu- 
lated through  an  equal  number  of  divisions  of  the  mine,  more 
or  less  equal  in  length,  with  volumes  of  greater  or  less  velocity, 
the  value  for  p  must  be  calculated  in  each  division  or  district 
separately;  and  for  each  differing  airway  the  aggregate  resist- 
ance in  each  division  is  the  sum  of  the  resistances  encountered 
in  each  of  its  various  galleries.  The  sum  of  the  several  fric- 
tional  losses  of  head  or  of  pressure,  and  that  pressure  or  head 
which  produces  the  final  velocity  at  the  mouth  of  the  mine,  is 
again  equal  to  the  ventilating  pressure  demanded  of  the  motor. 

Formerly  it  was  the  practice  to  meander  the  air  through 
all  the  galleries  of  each  lift  before  expelling  it  (Fig.  163).  This 
involved  heavy  pressures,  enormous  airways,  or  a  velocity  dan- 
gerously fast,  and  the  last  gang,  fed  by  the  departing  current, 
would  receive  an  irrespirable  atmosphere,  vitiated  by  the  emana- 
tions from  all  previous  sources.  There  was  nothing  to  commend 
this  pernicious  system,  and  it  is  certainly  a  matter  of  congratula- 
tion that  it  has  become  obsolete. 

Multiple  Air-circuits. — Mr.  J.  Buddie  introduced  a  system  of 
ventilation  for  fiery  mines  that  has  everything  in  its  favor.  This 
system  was  known  at  first  as  "coursing  the  air,"  and  now  is 
termed  "splitting  the  air,"  the  inception  of  which  is  due  to  Carlisle 
Spedding  or  his  son,  of  Whitehaven,  who  introduced  it  in  1 763.  By 
it  the  aggregate  quantity  of  air  is  increased,  the  dangers  of  explo- 
sion are  lessened  by  confining  its  train  of  evils  to  one  portion  of 
the  mine,  and  power  for  ventilation  and  haulage  is  saved,  since 
it  goes  hand  in  hand  with  the  method  of  panel- working  (Fig.  14). 
Each  panel  of  the  mine  is  completely  isolated  from  the"  contiguous 
districts  by  barrier  pillars,  and  is  ventilated  separately  by  delivery 
to  it  of  a  portion  of  the  volume  of  the  intake  which  does  service 
in  that  panel,  to  be  afterwards  discharged  into  the  return  air- 
way, where  it  rejoins  the  exhaust  from  the  other  districts.  The 
electric  distribution  for  purposes  of  illumination  and  the  water- 
supply  of  a  town  are  conducted  on  identically  the  same  prin- 


448  MANUAL  OF  MINIXG. 

ciple,  i.e.,  that  which  recognizes  the  tendency  of  a  fluid  to  seek 
a  shorter  and  easier  escape  from  confinement.  With  a  number 
of  conduits  receiving  at  a  common  point  a  volume  of  fluid  from 
a  larger  conductor,  each  will  convey  a  fractional  amount  of  that 
original  bulk  which  is  inversely  proportional  to  the  resistance 
offered  by  its  entire  rubbing-surface.  If  the  several  conduits 
again  meet  to  discharge  their  individual  volumes  of  fluid  at  a 
common  point  into  a  common  reservoir,  the  pressure  at  the  point 
of  discharge  is  the  same  at  the  mouth  of  each  and  every  pipe. 
Likewise  the  pressure  at  the  point  of  union  is  the  same  in  the  ends 
of  each  and  every  pipe.  The  loss  of  head  or  of  pressure  due  to 
the  flow  of  the  given  quantity  of  fluid  through  each  conduit  is 
then  the  same.  If  the  original  bulk  is  allowed  naturally  to  sub- 
divide, the  amount  of  fluid  in  the  several  branches  will  vary  in  an 
inverse  ratio  with  the  cross-sectional  area  of  their  conduits.  This 
is  equally  true  of  the  circulation  of  air  through  mine-galleries  or 
districts,  of  water  through  branching  pipes,  or  of  electricity 
through  connecting  wires  in  the  circuit. 

The  Process  of  Balancing  the  Resistances. — In  planning  the 
ventilating  system  for  a  mine  which  is  to  be  divided  into  a 
number  of  districts  for  ventilation  purposes,  the  practice  is  to 
calculate  for  each  separate  district  its  aggregate  resistance  to 
the  flow  of  the  volume  required  for  a  known  number  of  men 
employed  there,  and  for  a  dilution  of  the  gases  evolved  in  that 
district.  Several  separate  values  for  p  are  thus  obtained.  But 
these  district  resistances  must  be  equalized  or  else  the  inlet- 
current  will  be  so  subdivided  at  the  point  of  distribution  that  the 
large  bulk  of  the  air  will  pass  through  that  district  which  offers 
the  least  resistance ;  while  to  that  district  offering  the  greater  resist- 
ance the  volume  there  circulating  will  be  small.  This  is  usually 
the  reverse  of  the  requirements;  for,  generally,  that  district  offer- 
ing the  smaller  resistance  is  the  shorter  one,  having  less  men  in  its 
circuit,  and  therefore  requiring  a  smaller  volume  of  air;  while 
that  district  presenting  the  greater  resistance  to  the  flow  of  the 
current  is  either  more  extensive,  has  a  greater  volume  of  goaves, 
or  contains  more  working  places,  and  hence  demands  a  very 


DISTRIBUTION   OF   THE  AIR.  449 

large  fractional  part  of  the  main  current.  In  order,  then,  to 
automatically  deliver  to  that  district  requiring  more  air,  which 
at  the  same  time  offers  a  greater  resistance,  the  area  of  the  con- 
duit throughout  its  course  or  the  area  of  the  orifice  at  the  central 
point  of  distribution  must  be  made  sufficiently  large  to  tempt 
through  it,  or  into  it,  the  requisite  amount  of  air,  leaving  to  the 
smaller  district,  which  requires  less  air,  an  orifice  of  entry  which 
is  comparatively  small.  By  so  doing  the  differences  in  pressure 
for  each  and  every  district,  between  the  point  at  which  the  splits  of 
the  fresh-air  current  are  made  and  the  point  at  which  the  return 
currents  from  the  same  splits  reunite,  can  be  equalized  to  that 
of  the  one  offering  the  greatest  resistance.  Then  the  current 
will  naturally  divide  according  to  the  areas  of  the  inlets  or  of 
the  passages,  and  each  district  will  receive  its  apportioned  frac- 
tion of  the  incoming  air.  Hence,  whenever  the  ventilation  of 
the  mine  is  to  be  split  into  several  currents  and  the  air  is  to  be 
apportioned  in  accordance  with  the  demand,  the  mine  foreman, 
having  calculated  the  relative  values  for  the  head  lost  in  each, 
determines  by  proportion,  as  will  be  seen,  the  comparative  area 
of  inlet  to  be  provided  the  several  districts  at  the  point  of  dis- 
tribution, and  there,  by  means  of  doors  and  other  regulators, 
furnish  to  the  given  district  the  area  desired. 

The  measurements  of  the  water-gauge  pressure,  or  loss  in  head 
between  the  beginning  and  the  end  of  the  split  and  the  velocity 
of  the  flow  of  air,  are  made  in  the  intake;  and  while  it  is  not 
always  possible  to  subdivide  the  current  at  a  common  point 
of  distribution,  this  should  be  done  as  near  to  the  downcast  as 
circumstances  will  admit.  The  same  may  be  said  of  the  point 
of  reunion.  Otherwise  the  resistance  of  the  intermediate  ways 
and  of  the  entries  must  be  determined  and  provided  for,  as  may 
be  seen  in  the  example  given  below.  The  aggregate  resistance 
of  the  intermediate  ways  of  the  several  districts  through  which 
the  air  is  circulated  determines  the  maximum  number  of  splits 
which  are  possible. 

Simple  as  is  the  theory,  and  satisfactory  and  economical 
as  is  the  plan  when  well  developed,  it  is  not  easy  of  execution- 


45°  MANUAL  OF  MINING. 

The  success  of  the  plan  involves  an  exact  manipulation  and 
great  skill  in  taking  due  precautions  to  balance  the  various  factors, 
to  determine  the  equilibrium  designed,  and  to  prevent  one  panel 
or  district  from  receiving  too  brisk  a  current  at  the  expense  of 
others.  Hence,  while  it  is  eminently  desirable  to  apply  this 
theoretical  distribution,  its  difficulty  is  recognized,  and  it  has 
become  the  practice  of  the  foreman  to  approximate  the  desired 
conditions  by  making  repeated  tests  upon  the  quantity  of  air 
flowing  in  a  given  circuit,  which,  if  insufficient,  is  provided  for 
by  enlarging  the  inlet  area  for  the  given  district  and  watching  its 
reaction  upon  the  volumes  in  the  other  dependent  splits.  In 
shallow  workings,  though  the  mine  may  be  extensive,  the  practice 
is  an  inexact  one  in  many  cases.  It  may  often  be  cheaper  to 
sink  a  new  shaft  to  furnish  separate  ventilation  to  a  district  than 
it  would  be  to  undertake  to  furnish  an  elaborate  system  of  splits. 
The  Power  Required  for  Ventilation. — Though  it  may  not 
require  a  demonstration  to  show  that  the  subdivided  splits  of 
the  current  are  productive  of  greater  economy  in  ventilating 
power,  attention  will  be  called  to  the  fact  that  the  ventilating 
force  in  horse-power,  necessary  to  deliver  a  volume,  Q,  against  a 
mine  resistance,  p,  is  measured  by  the  expression 

H.P. = Q/»-J- 33,000 = Qm-j- 6365.7. 

The  resistance  which  would  be  offered  by  the  aggregate 
of  all  the  districts  to  the  flow  of  the  entire  volume,  Q,  through 
the  whole  length  of  the  circuit  is  measured  by  p.  Each  frac- 
tional volume,  q,  qf,  q',  etc.,  passing  through  only  one  branch 
of  the  circuit  would  offer  a  resistance  r,  rf,  r" ,  etc.,  which  is  very 
small  compared  with  p.  When  the  mine  boss  has  adjusted  the 
regulator  doors  at  the  point  of  distribution  by  altering  the  respec- 
tive areas  of  inlets,  the  resistances  in  all  of  the  several  circuits 
are  equalized,  the  work  performed  in  each  split  is  qr,  qfr,  q"ry  etc., 
and  the  aggregate  ventilating  power  is  their  sum.  As 

q+qf+q"+.    .    .  =Q    and    r<p, 


DISTRIBUTION   OF  THE  AIR.  451 

the  power  required  for  the  ventilation  in  branches  will  be  less 
than  that  for  a  single  current,  (),  through  the  same  passages. 

The  power  required  for  16,200  cubic  feet  of  air  flowing  in  one 
column  would  be  capable  of  producing  70,884  cubic  feet  of  air  in 
five  splits,  94,850  cubic  feet  in  ten  splits,  and  nearly  100,000  cubic 
feet  in  fifteen  splits. 

EXAMPLES. — i.  A  colliery  is  ventilated  by  a  Guibal  fan  of  21  feet  3  inches 
diameter,  making  40  rev&lutions  per  minute.  How  many  cubic  feet  will  it 
produce?  The  air  must  pass  through  a  main  airway  300  feet  long,  6'Xi2', 
before  being  split  up  into  three  separate  airways,  one  being  12,000  feet  long, 
6'Xs';  another,  n,:oo  feet,  of  area  6X7';  while  the  third  is  10,000  feet 
long  and  5'X8'  in  section.  Required  also  the  water-gauge  pressure,  assum- 
ing the  two  shafts  together  to  consume  0.226  Ib.  per  square  foot  in  .riction. 

16,380  cubic  feet  and  0.9  inch. 
The  splits  are  all  drawn  from  a  common  point  of  junction. 

Theoretically,  the  fan  produces  a  water-gauge  pressure  of  0.902  inch 
(4.677  Ibs.).  Then  the  entire  mine  offers  a  resistance  of  3.504  Ibs.  (0.676 
inch).  The  resistance  of  the  main  airway  is  p  —  0.00000000062 79(2 2.  Q, 
the  quantity  of  air  delivered,  is  divided  up  into  three  separate  sections,  accord- 
ing to  their  resistances.  As  p  is  the  same  for  each,  the  quantities  q  may  be 
known  in  terms  of  p  to  be 

2i7iV7,     3455^,    and    3368\/^ 

Now  the  resistance  of  the  entire  mine  is  equal  to  3.504  Ibs.,  plus  that  of 
the  main  airway,  plus  that  of  the  split  ,  p.  From  this  we  know  ^=3.504 
— 0.00000000062 79(8994^) 2,  whence  p  becomes  3.333  and  the  r  sistance  of 
the  main  airway  0.170  Ib.  Q  then  becomes  16,380  cubic  feet,  and  the  quanti- 
ties received  by  the  three  splits  are  3940,  6270,  and  6110  cubic  feet.  (The 
difference  in  results  arises  from  failing  to  carry  out  the  decimals  beyond  two 
places.) 

2.  An  airway  3000  feet  long,  8'X4'  area,  is  carrying  20,000  cubic  feet. 
How  many  cubic  feet  would  be  delivered  if  the  air  was  split  into  three  currents, 
the  power  remaining  the  same?  The  sections  are  3000  feet  long  and  8'X  4' 
area;  3600  feet  and  s'Xg';  and  4800  feet  6'Xio'. 

51,736  cubic  feet  and  11.56  H.P. 

The  calculated  power  necessary  to  drive  the  quantity  of  air,  Q,  through  the 
three  sections  is  equal  to  the  sum  of  the  three  powers,  pav,  of  each  section. 

The  benefits  that  may  be  derived  from  splitting  the  air- 
current  are  manifest  by  inspection  of  the  following  case: 


45  2  MANUAL  OF   MINING. 

EXAMPLES.  —  i.  A  mine  has  two  slope  entries,  9'Xi4'  in  cross-section  and 
100  feet  long,  and  such  internal  resistances  as  would  be  equivalent  to  8000  feet 
of  a  typical  airway  5'Xio'  in  cross-section.  What  pressure  and  power  will 
be  requisite  to  propel  16,200  cubic  feet? 

^=0.2619  and  u=4243  ft.-lbs.,  for  the  two  entries,  and 
^=11.19  and  «=  181,383  for  the  total. 

2.  Required  the  quantities  of  air  that  will  circulate  where   there  are 
2,  3,  5,  10,  and  15  equal  splits,  the  pressure  remaining  the  same  as  above. 

After  calculating  the  pressure  p  for  the  one  current  as  above,  then  proceed 
to  ascertain  the  pressure  p'  necessary  to  circulate  16,200  cubic  feet  in  the 
several  cases.  These  will  be  found  1.369,  0.405,  0.0874,  0.0109,  and  0.0032 
Ib.  per  square  foot,  respectively;  as  the  area  of  each  airway  is  50  sq.  ft.,  the 
aggregate  for  the  equal  splits  are  100,  150,  250,  500,  and  750  square  feet  in 
the  several  cases,  while  .he  rubbing-surfaces,  Im,  are  the  same  (240,000 
quare  feet).  The  pressures  are  then  apportioned  directly  to  10,935  Ibs., 
the  mine  friction  of  one  current  as  above.  Whence,  the  pressures  being  as 
the  squares  of  the  volumes  circulating,  we  obtain  33,409  cubic  feet,  66,354 
cubic  feet,  91,692  cubic  feet,  103,755  cubic  feet,  and  105,255  cubic  feet, 
and  373,887  ft.-lbs.,  742,504  ft.-lbs.,  1,026,030  ft.-lbs.,  i,  161,010  ft.-lbs., 
and  1,177,760  ft.-lbs.  as  the  respective  powers  «. 

3.  If  it  be  desired  to  know  what  quantities  will  circulate  with  the  same 
power  M,  as  in  Ex.  2,  then  we  have  but  to  apportion  the  volumes  to  the 
cube  roots  of  the  powers,  «. 


Thus    ^373,887    :  33,409:  :i8i,383  126,252,  the  volume  with     2  splits. 

^742,504    :  66,754:  :3/iSi,  383  :  41,  479.    "        "          "  3       " 

"ii/  1,  026,030:  91,692:  :^//i8i,383:5i,46i,    "        "          "  5       " 

3/i,  161,010:  103,  55:  :3/  181,383:  55,881,    "        "          "  10      " 

^1,177,  760:  105,255  ::^/i8i,383:  56,419,    "        "          "  15      " 

4.  When,  however,  he  splits  are  not  taken  from  a  common  point  of 
functure,  the  procedure  for  ascertaining  the  mine  resistances,  and,  subse- 
quently, for  balancing  the  delivery  of  air  to  the  several  sections,  is  not  so 
simple.  The  plan  consists  in  determining  the  several  resistances  and  the 
powe-  necessary  to  overcome  them.  These  are  then  added  as  fo'lows:  In 
Fig.  163  D  is  the  downcast  shaft,  84^  feet  deep,  8'Xio'  in  cross-section, 
delivering  56,000  cubic  feet  per  minute;  by  force-fan,  27,750  cubic  feet  go  to 
the  left  gangway,  while  the  right  gangwav  passes  28.250  cubic  feet  of  a?r. 
The  water-gauge  stands  at  f  inch  (4.525  Ibs.).  Required  the  volumes  of  air 
received  by  the  splits  A,  B,  C,  and  D. 


DISTRIBUTION   OF   THE  AIR.  453 

The  distances  along  the  gangway  at  which  the  splits  are  taken  are,  a,  460 
feet  from  d;  b,  960  feet;  and  c,  1360  feet  from  d.  Dimensions  of  gangway 
6'Xi2'. 

The  splits  are  E,  receiving  5000  cubic  feet  through  100  feet  of  6'Xia' 
gangway;  60  feet  of  4'X2f  break-through;  and  60  feet  of  return  airway 


c  b  a 

FIG.  163. — Diagram  of  a  Split  Air-current. 

5^X14';  C,  having  a  resistance  equivalent  to  2401  feet  of  typical  roadway 
15' X  6.87',  connecting  with  the  return-way  at  cv  941  feet  from  the  upcast  U; 
B,  which  has  220  feet  of  gateway  4'X;',  1400  feet  of  entry  4'X8',  and  1000 
feet  of  room  4/Xi2/,  delivering  at  bv  a  point  581  feet  from  U;  and  A,  which 
has  a  resistance  equivalent  to  1810  feet  of  typical  roadway  5'Xio',  connect- 
ing with  the  return  airway  313  feet  from  U. 

The  upcast  shaft  is  870  feet  deep  and  12  feet  in  diameter.  It  was  found 
that  the  velocity  of  the  outgoing  air  was  nearly  708  feet.  This  would  cor- 
respond to  about  80,000  cubic  feet  of  air. 

In  the  return  airway,  along  ual,  the  volume  was  found  to  be  41,610  feet, 
This  shows  an  increase  in  volume  of  13.870  cubic  feet,  which  may  be  accounted 
for  by  the  higher  temperature  of  the  outtake  air  or  increments  to  the  original 
volume  from  gas. 

Beginning  at  E,  where  5000  cubic  fee.  are  circulating,  we  find  the  total 
resistance  to  b  such  as  requires  a  pressure  of  0.096  Ib.  per  square  foot  at  c, 
to  overcome  it.  Hence  the  same  pressure  must  exist  at  the  mouth  of  C, 
wher?  pa= 0.000000219677"  Equating  the  two  values  for  p  (q  being  unknown 
for  C,  and  5000  cubic  feet  for  E)  we  find  that  the  volume  of  air  of  6410  cubic 
feet  offers  a  proportionate  resistance  to  that  in  E. 

At  the  point  b  the  total  resistances  of  E,  C,  and  their  gangways  must  be 
proportional  to  those  of  B,  a,-*,  the  relative  volumes  to  be  circulated.  At  b  the 
11,410  cubic  feet  passing  out  produces  a  pressure  of  0.57  Ib.  per  square.foot,  at 
which  rate,  by  equating  this  with  the  sum  of  the  resistances  of  the  split  B,  we 
find  that  q  becomes  3720  cubic  feet. 

Tn  like  manner  12.580  cubic  feet  will  be  found  Capable  of  passing  through 
A,  the  frictional  resistance  at  a  being  1.19  Ibs.  per  square  foot. 

The  total  resistances  pa  of  the  various  sections  are:  da,  53.33  Ibs.  •  A,  74.59 
Ibs.;  ab,  17.29  Ibs.;  B,  16.04  Ibs.;  be,  7.83  Ibs.;  C,  9  Ibs.;  E,  6.02  Ibs.;  c,6,, 
17.67  Ibs.;  fejflj,  23.29  Ibs.;  a,«,  91.2  Ibs.;  D,  323.85  Ibs.,  and  U,  356.2  Ibs. 


454  MANUAL  OF  MINING. 

The  volumes  of  air  are,  respectively,  27,740  cubic  feet;  12,580  cubic  feet; 
15,150  cubic  feet;  3720  cubic  feet;  11,410  cubic  feet;  5910  cubic  feet;  5000 
cubic  feet;  17,115  cubic  feet;  22,725  cubic  feet;  41,610  cubic  feet;  80,000 
cubic  feet;  and  56,000  cubic  feet. 

The  pressure  mentioned  is  only  that  necessary  to  overcome  the  rubbing 
friction  of  the  moving  air  against  the  sides.  No  cognizance  has  been  taken  of 
the  pressure  requisite  to  air  the  mine.  That  power  will  depend  upon  the 
velocity,  and  is  equal  to  about  0.17  Ib.  pressure  per  foot. 

The  power  requisite  to  force  the  air  into  and  out  of  the  mine  (11.9  Ibs.  per 
square  foot)  is  666,400  ft. -Ibs. 

5.  An  airway  5'X8'  across,  6000  feet  long,  is  followed  by  a  through 
2'X5'  in  cross-section  300  feet  long.  To  drive  10,000  cubic  feet  of  air  through 
them  requires  5.5  horse-power.  Their  frictional  resistances  are  equivalent 
to  5.2893  Ibs.  and  9.114  Ibs.  per  square  foot,  respectively;  the  total  resistances, 
pa,  being  211.57  and  91.14  Ibs.  A  pressure  of  302.71  Ibs.  is  therefore  requisite 
at  the  inlet  to  overcome  frictions  of  the  passage.  302.71  -^40 X  10,000=  75,667 
ft. -Ibs.  to  propel  the  air  through  the  first  section,  in  addition  to  which  it  requires 
9.114X10,000  Ibs.  to  overcome  the  resistance  of  the  second. 

When  the  "through"  precedes  the  airway,  we  have  no  differ  nee  in  fric- 
tional resistances,  but  the  total  pressure  in  the  first  section  must  be  such  as  to 
overcome  the  total  resistance  of  the  two,  hence  it  is  302.71  Ibs.  (30.27  Ibs.  per 
square  foot).  Then  30.27X10,000  +  5.289X10,000=355.601  ft.-lbs.=  10.7 
horse-power. 

If  the  "through"  is  midway  along  the  6ooo-foot  airway  and  has  300  feet 
of  the  5' X  8'  roadway  on  each  side,  we  still  have  the  same  resistances,  yet  the 
power  requisite  for  propulsion  is  299,047  ft.-lbs.  For  each  end  the  friction  is 
2.645  Ibs.  per  square  foot,  and  for  the  middle  it  is  still  9.114  Ibs.  per  square 
foot.  The  total  pressure  of  th?  first  is  105.784  Ibs.;  of  the  second,  91.14  Ibs.; 
and  of  the  last,  105.784  Ibs.  The  pressure  at  the  inlet  is  302.71  Ibs.  (7.5677 
Ibs.  per  square  foot);  at  the  entrance  to  the  second,  196.924  Ibs.  (19.6924  Ibs. 
per  square  foot);  and  at  the  entrance  to  the  last,  105.784  Ibs.  (2.645  Ibs. 
per  square  foot). 

Goaves. — These  are  the  most  dangerous  places.  Here  the 
waste  "rock,  broken  coal,  slate,  dust,  pyrites,  etc.,  accumulate  in 
the  presence  of  water  which,  with  the  stagnant  air,  induce  a  com- 
bustion from  which  sulphuretted  hydrogen  and  other  gaseous 
products  exude.  It  is  estimated  that  the  air-space  in  a  goaf  is  one 
sixth  of  the  volume  of  coal  extracted,  and  in  it  most  likely  will 
breed  a  great  deal  of  gas,  of  which  the  sweating  of  the  roof  is 
an  infallible  sign.  Often  a  water-gauge  placed  in  a  goaf-stopping 


DISTRIBUTION   OF   THE  AIR.  455 

will  indicate  by  the  difference  of  level  in  its  arms  whether  or  not 
any  accumulation  of  gas  exists  behind  the  stopping.  Spontaneous 
combustion  once  begun  therein,  nothing  will  stop  it.  For  this 
reason,  though  aeration  is  possible,  fiery  coals  should  only  be 
worked  by  a  method  involving  complete  removal  of  the  coal,  or 
its  replacement  by  clean  waste. 

The  Velocity  of  the  Air-current. — Spacious  air-drifts  will 
then  reduce  the  resistances  and  also  the  velocity  of  flow.  It  is 
not  desirable  that  the  velocity  exceed  500  feet  per  minute,  nor  is 
it  comfortable  or  safe.  A  speed  exceeding  500  feet  is  equally 
injurious  with  stagnation.  Several  mining  commissions  of 
Europe  have  experimentally  determined  that,  aside  from  the 
chilling  effect  of  walking  against  so  rapid  and  cool  a  breeze, 
lamps  are  not  safe:  a  rapid  current  incites  explosion  by  driving 
the  gases  through,  or  the  flame  against,  the  screens.  Many 
lamps  can  resist  higher-speed  currents,  but  none  are  safe  in 
over  900  feet  per  minute.  In  the  old  country  the  air-currents 
at  different  parts  of  the  mines  vary  in  velocity — at  the  coal  face 
often  as  fast  as  900  feet;  but  here  our  Davy,  Stephenson,  or 
Clanny  lamps  require  protection  in  such  a  velocity,  which  exceeds 
that  of  American  practice. 

The  Anemometer. — A  miner  can  approximately  estimate  the 
speed  of  the  current  by  knowing  the  rate  at  which  he  must  walk 
to  keep  the  flame  erect,  or  by  noting  the  time  elapsing  between 
the  discharge  of  a  volatile  fluid  or  smoke  and  the  time  of  its 
arrival  at  a  point  a  known  distance  beyond.  The  anemometer 
is,  however,  much  the  simpler  instrument  for  measuring  the 
velocity  of  the  flowing  air.  In  a  case  is  a  series  of  vanes  which 
are  moved  by  the  current,  and  these  by  proper  gearing  turn 
indices  over  their  respective  dials  at  such  a  rate  that  the  velocity 
may  be  at  once  read.  It  does  not  give  accurate  results  on  ac- 
count of  the  friction  of  its  mechanism.  Each  instrument  has 
its  own  factor,  which  is  not  even  constant.  Biram's  and  Cos- 
tello's  patterns  (Figs.  164  and  165)  are  most  used  in  America. 
Their  factors  are  ascertained  by  occasional  test,  the  anemom- 


456 


MANUAL   OF  MINING. 


eter  being   revolved  on  a  whirling   table,  and   its  reading  com- 
pared with  the  actual  velocity  of  revolution. 

The  point  selected  for  observing  the  velocity  should  be  in  a 
straight  gallery,  whose  sides  and  roof  are  a  fair  average  in  rough- 
FIG.  164.  FIG.  165. 


Anemometers. 

ness,  and  where  there  is  neither  a  sudden  bulge  nor  a  contraction. 
The  average  of  several  one-minute  readings  is  taken  at  the  place 
of  measurement,  near  the  roof,  sides,  and  centrally.  Then  the 
cross-section  of  the  conduit  at  the  observing  station  is  taken. 

The  Efficiency  of  a  Ventilating  System. — One  object  in  deter- 
mining the  velocity  of  the  air  circulating  in  the  passage  is  to 
ascertain  the  efficiency  of  the  motors  supplying  the  current. 
For  this  purpose  the  cross- sectional  area  of  the  airway  at  the 
observing  station  is  measured,  and  the  product  of  this  with  the 
velocity  and  the  ventilating  pressure  at  the  given  point  deter- 
mines the  actual  work  of  the  effective  power  in  the  air  at  that 
point.  The  power  U  to  move  the  air  is  o.oooo303-yaP.  The 
ratio  between  the  result  of  these  observations  and  the  simultane- 
ous indicator  diagrams  of  the  engine  gives  the  mechanical  effi- 
ciency of  the  ventilating  system.  The  ultimate  comparisons  of 
this  work  in  terms  of  consumption  of  coal  should  show  for  a 
good  system  not  more  than  n  Ibs.  of  fuel  per  useful  horse-power 


DISTRIBUTION  OF  THE  AIR.  457 

produced  by  the  fan  and  40  Ibs.  to  70  Ibs.  per  effective  horse- 
power of  the  furnace. 

The  Regulation  of  the  Multiple  Circuits. — The  air  entering 
the  mine  is  conducted  to  the  men  by  a  route  as  direct  as  the 
plan  of  the  workings  admits.  To  those  who  are  at  work  in 
the  shaft  and  at  the  breast  of  the  exploration  entries,  the  current 
is  led  by  a  partition  in  the  shaft  or  drift-forming  conduits  for 
an  intake  and  return  current;  to  the  men  engaged  in  extract- 
ing coal,  the  current  passes  from  the  main  entry  to  the  head- 
ing, thence  through  cross-headings,  or  branch  headings,  to  the 
rooms.  The  air-current  passing  down  the  entries  by  either 
shaft  or  slope  is  split  at  each  level  or  lift  to  a  right  and  a  left 
branch.  In  each  lift  the  current  flows  to  the  furthermost  point, 
along  the  main  heading,  thence  rising  into  the  most  remote 
room,  along  its  face,  thence  through  ''dog-holes,"  progressing 
toward  the  hoistway  till  each  room  has  been  traversed,  when  the 
current  is  led  to  the  return  or  back  heading  which  communi- 
cates to  the  upcast  way,  there  to  be  joined  by  the  return  current 
from  the  opposite  side  of  the  lift,  and  finally  to  unite  with  the 
return  current  rising  from  the  lower  lifts  which  have  been  fed  in 
like  manner. 

The  main  heading  is  usually  the  haulage-way,  the  back  head- 
ing serving  only  for  the  vitiated  air.  It  may  be  used  for  a  travel- 
ling way.  The  jaw  of  each  room,  except  that  one  which  is  most 
advanced,  is  stopped  up,  which  latter,  in  time,  is  closed  as  an- 
other room  is  turned  from  the  heading.  As  each  room  must  be 
in  free  communication  with  the  haulage-way  for  the  delivery  of 
its  coal,  the  stopping  there  provided  must  be  capable  of  opening 
for  the  passage  of  the  coal.  The  swinging  doors,  therefore, 
which  are  placed  in  the  entry  are  always  closed  to  prevent  the 
air  from  returning  at  once  down  the  back  heading.  Here  doors 
are  built  to  deflect  it  into  the  rooms,  with  also  a  brattice  in  each 
room,  to  direct  the  current  it  receives  from  its  neighbor.  From 
the  last  room,  whether  working  or  abandoned,  the  air  passes  to 
the  return  current.  In  the  lower  portions  of  the  room  not  in 
the  direct  sweep  of  the  air-current  the  air  is  prevented  from 


4S8  MANUAL  OF  MINING. 

becoming  stagnant  by  the  leakage  from  the  main  current  and 
the  addition  of  such  volume  as  is  swept  in  or  out  by  the  travel  of 
the  cars  from  and  to  the  haulage-way. 

The  distribution  in  the  long  wall  is  perfectly  similar,  in  which 
case  the  gob  roads  are  stopped  up  by  doors  to  convey  the  current 
to  the  extremities  of  the  workings,  whence  the  air  flows  along 
the  long- wall  face. 

In  flat  seams,  the  double  entries  of  which  have  a  low  grade 
for  haulage  from  the  surface,  the  similar  split  system  is  em- 
ployed, the  splits  being  taken  at  the  point  of  union  with  the  main 
headings.  When,  by  reason  of  the  low  pitch,  three  entries  are 
driven  from  daylight,  the  common  practice  consists  in  making  the 
central  entry  the  return  airway,  the  air  being  led  into  the  mine 
in  two  currents  on  either  side.  This  simplifies  the  subdivision 
of  the  currents,  which  may  be  effected  with  few  doors.  Each 
district  on  either  side  will  have  but  one  crossing  of  its  return 
current  with  the  intake  current. 

Structural  Appliances  Required  for  Controlling  the  Air- 
current.— Their  object  is  to  confine  and  direct  the  air-currents 
in  a  given  path.  They  may  be  permanent  or  temporary  in  char- 
acter; they  may  also  serve  as  stoppings  to  prevent  any  circula- 
tion through  a  given  way,  or  as  a  means  of  deflecting  the  current 
into  some  other  channel. 

In  the  split  system  of  ventilation  several  small  currents  radi- 
ating from  one  given  centre  of  distribution,  or  branching  from 
a  main  airway,  are  returned,  after  having  done  service,  to  an- 
other airway,  whence  they  are  carried  to  the  surface.  Provision 
must  be  made  for  carrying  the  return  current  over  or  under  the 
intake,  where  they  intersect  for  regulating  the  amount  of  air 
delivered  to  each  branch  of  the  mine;  and  for  flexible  stop- 
pings along  the  gangway,  admitting  of  the  passage  of  trains  of 
cars. 

The  Means  for  Deflecting  the  Current. — These  are  of  seven 
varieties: 

Air-bridges,  constructed  to  carry  the  return  current  over  01 
under  the  airway. 


DISTRIBUTION   OF   THiL  A.ii.  459 

Stoppings,  or  walls,  closing  the  "throughs"  when  their  ser- 
vice as  connectors  has  been  performed. 

Partitions  of  wood  or  canvas  to  temporarily  divide  each  air- 
way into  two  compartments  for  an  inlet  to  the  face  and  an  out- 
let to  the  cross-heading  nearest  to  the  face,  which  in  turn  con- 
nects with  the  return  air,  back  heading,  or  adjoining  room. 

Brattices  of  canvas  hung  from  the  roof  in  the  common 
travelling-way  to  serve  as  a  light  temporary  resistance  to  the 
flow  of  air,  without  impeding  the  access  of  the  men  into  the 
rooms. 

Swinging  doors  in  the  gob  roads  and  at  the  jaws  of  the  rooms 
to  prevent  the  entry  of  air  from  the  main  heading,  but  to  allow 
admission  to  the  men. 

Swinging  doors  at  the  return  heading  at  appropriate  places 
to  check  the  immediate  return  of  the  current  and  to  deflect  it 
to  a  room. 

Sliding  doors  placed  in  the  branch  airways  and  fixed  so  as  to 
offer  an  area  for  the  inlets  to  several  branches,  including  the  con- 
tinuation of  the  mainway. 

The  Stoppings. — These  consist  of  walls  strongly  built  to 
resist  the  pressure  of  the  roof  and  floor  and  also  to  prevent  the 
leakage  of  air.  They  are  of  brick,  with  their  faces  set  a  little 
back  from  the  wall  of  the  gallery  in  order  to  provide  for  any 
peeling  of  the  coal  from  the  side  of  the  airways.  These  walls  are 
built  on  either  side  of  the  airway  two  or  three  bricks  thick,  and 
often  filled  with  rubble  to  make  a  firm  pillar  and  support.  Oc- 
casionally their  exposed  faces  are  plastered  with  coal-tar  or 
cement  to  make  them  impervious  They  are  built  about  the 
goaves  and  other  gaseous  districts. 

The  temporary  stoppings  are  of  wood,  or  canvas  sheets,  more 
or  less  firmly  framed,  and  may  be  movable,  as  are  doors,  or  fixed, 
as  in  the  case  with  curtains  and  panels. 

Air-crossings  or  Bridges. — These  are  placed  at  the  intersec- 
tion of  each  pair  of  cross-entries  to  allow  the  pure  air  from  the 
intake  to  enter  the  mine  without  interference  from  the  vitiated 
air  in  its  exit.  These  bridges  are  carried  over  the  roadways  and 


460  MANUAL  OF  MINING. 

dispense  with  the  use  of  doors.     They  are  permanent  structures 
whose  position  is  determined  by  the  plan  of  the  mine. 

Overcast. — There  are  two  classes  of  air-bridges,  one  in  which 
the  return  airway  is  carried  over  the  main  haulage- way  and  intake, 
and  the  other  in  which  it  may  pass  below  the  intake.  Though 
there  are  many  conditions  favoring  the  undercast  crossing  the  air- 
bridge prevails  (Fig.  166).  In  this  case,  at  some  distance  before 
the  point  of  intersection  of  the  two  airways  is  reached,  an  excava- 
tion is  made  in  the  roof  sufficiently  high  and  wide  to  afford  an 
airway  of  ample  dimensions,  which  excavation  is  carried  above 
the  intake  airway  the  desired  distance  and  down  on  the  other 
side  to  the  main  level  of  the  coal-seam  at  some  distance  beyond. 
The  coal  close  to  the  sides  of  the  intake  airway  is  left  undis- 
turbed. A  bridge  is  then  built  consisting  of  a  heavy  flooring  of 
timber  laid  above  the  opening  to  make  a  roof  for  the  intake  and 
a  floor  for  the  intake  airway.  It  would  be  cheaper  to  remove 
all  the  coal  on  the  sides  and  erect  stopping-walls  to  furnish  sup- 
port for  the  bridge  overhead,  but  these  are  not  always  air-tight. 
It  is  absolutely  essential  that  the  partitions  between  the  return 
and  the  intake  airway  should  be  air-tight.  The  cost  of  an  over- 


FIG.  1 66. — A  Rock-tunnel  Air -crossing. 

cast  is  about  $125.00,  with  a  month's  labor  of  common  help 
and  ten  days  of  two  masons.  There  would  be  required  perhaps 
100  bushels  of  mortar,  5000  bricks,  and  6  cu.  yds.  of  sand. 

Many   engineers   prefer  to   drive   the  overcast   through   the 
solid  rock  to  insure  a  tight  airway  and  one  which  is  not  liable 


DISTRIBUTION   OF   THE  AIR. 


461 


to  be  influenced  or  affected  by  the  accidents  in  the  intake.    These 
are  more  expensive  to  build,  but  have  many  advantages. 

The  Undercast. — This  air- crossing  is  built  below  the  floor  (Fig. 
167)  of  requisite  dimensions  and  far  enough  beyond  the  intake  on 


FIG.  167. — An  Undercast  Air-crossing. 

either  side  to  leave  a  secure  and  undisturbed  pillar  for  a  stopping 
in  the  latter.  It  is  framed  and  boxed  as  is  the  overcast.  In  a 
few  cases  this  construction  may  be  cheaper  than  the  overcast, 
and  is  preferred  when  the  roof  cannot  well  be  disturbed.  But 
the  liability  to  undermine  the  coal-pillars  which  rest  upon  the 
floor  of  the  seam  and  the  risk  of  creep  constitute  a  serious  objec- 
tion to  it.  The  depression  which  is  made  in  the  floor  furnishes 
opportunity  for  lodgement  of  water  and  mine-gas  and  may  be- 
come in  time  absolutely  impassable.  It  is  always  subject  to 
constant  flooding. 

How  Should  Air-bridges  be  Made? — It  is  a  disputed  question 
whether  air-bridges  should  be  carried  through  the  solid  rock 
or  be  built  in  the  roof.  Both  serve  equally  well  their  pur- 
poses under  general  conditions.  But  the  effect  of  an  explosion 
upon  the  two  is  different.  The  rock  terminal  airway  remains 
intact  and  the  ventilation  of  the  mine  is  unimpaired.  The 
boxed  bridge- crossing  is  usually  destroyed  by  the  force  of  the 
blast,  and  with  it  the  air-current.  Of  course  the  force  of  the 
blast  is  confined  to  the  airways,  and  its  violence  may  produce 
heavy  falls  which  will  afterwards  present  obstacles  to  the  rescuing 
party,  when  the  work  of  rehabilitation  begins,  though  the  number 


462  MANUAL  OF  MINING. 

of  such  violent  explosions  that  will  shatter  the  roof  will  be 
small. 

The  timber  overcast  bridge  furnishes  a  safety-valve  for  the 
force  of  the  explosion  and  prevents  much  damage.  In  such 
event  the  reestablishment  of  the  air-current  after  an  explosion 
becomes  a  difficult  matter.  The  wooden  partitions  may  be 
destroyed  by  a  moderate  explosion,  and  the  number  of  occasions 
when  the  rescuing  party  must  build  bridges  and  stoppings  is 
larger  than  those  in  which  it  must  reopen  a  blocked  airway 
through  a  solid  overcast. 

A  trap-door  in  the  floor  of  an  overcast  which  would  raise  with 
the  explosion  and  fall  again  with  the  heavy  reaction  that  follows 
would  enable  the  air-current  to  be  reestablished  and  meet  the 
objection  to  the  air-bridge.  Such  a  door  could  be  built  cheaply 
and  can  be  made  tight  for  ordinary  emergencies. 

The  greater  expense  of  the  rock  tunnel  would  result  in  a 
smaller  number  of  them  being  driven,  and  hence  less  split  air 
currents.  As  the  economic  distribution  of  air  depends  upon 
the  use  of  numerous  air-currents,  there  would  be  a  greater  econ- 
omy secured  with  the  same  outlay  by  an  abundant  use  of  the 
cheaper  air-bridges  than  would  be  obtained  for  a  similar  number 
of  solid  overcasts. 

The  mining  laws  of  Pennsylvania  require  overcasts  and  air- 
bridges of  masonry  or  in  solid  rock,  in  all  mines  generating  fire- 
damp in  sufficient  quantities  to  be  detectible  by  the  ordinary 
safety-lamps. 

An  air-bridge  passing  over  a  slope  is  always  built  in  the  solid. 

Mine-doors. — The  doors  are  of  two  classes — those  provided 
with  an  additional  valve  or  gate,  which  may  be  opened  or  closed 
within  a  limited  range;  and  those  doors  swinging  on  hinges, 
which  may  be  opened  for  purposes  of  ingress  or  egress;  they 
should  open  against  the  direction  of  a  possible  inlet  current, 
in  order  to  completely  direct  the  current  along  the  main  way. 
Doors  are  the  main  dependence  for  the  ventilation  of  mines,, 
particularly  those  which  are  subdivided  into  districts  for  pur- 
poses of  ventilation.  They  are  placed  in  such  position  as  will 


DISTRIBUTION   OF   THE  AIR.  463 

temporarily  and  effectually  check  and  deflect  the  current. 
Wherever  their  location,  they  must  be  built  with  great  care,  of 
matched  timber,  closely  fitted  in  their  frame,  and  are  maintained 
only  so  long  as  the  draught  is  to  be  kept  up.  Some  are  even 
provided  with  weather-strips  on  their  edges.  They  are  placed 
in  all  the  side  entries  to  the  rooms,  swinging  outward  towards  the 
main  passageway,  and  are  generally  in  pairs  when  located  in 
the  haulage-way.  In  the  latter  case  they  are  located  far 
enough  apart  that  the  two  will  not  be  open  at  the  same  time, 
and  thus  interrupt  the  principal  circulation. 

Regulator-doors  are  employed  (Fig.  168)  at  the  mouth  of  entry 
to  a  ventilating  district  in  the  mine,  the  gate  of  which  constitutes 
an  adjustable  sliding  door  capable  of  being  secured  against 
disturbances,  and  are  employed  to  regulate  the  supply  of  air 
to  be  delivered  to  that  district  in  accordance  with  the  system 
described  above.  The  placing  of  the  slides  is  left  to  the  mine- 
boss  or  the  fire-boss.  Those  doors  which  are  placed  in  haulage- 
ways  either  require  an  attendant  or  have  their  frames  so  inclined 
that  the  door  swings  to  of  itself.  Automatic  mechanical  appli- 
ances for  opening  and  closing  the  doors  from  a  distance  without 
stopping  the  mules  or  other  haulage-motors  are  employed  in 
many  collieries,  but  do  not  stand  in  great  favor.  While  there 
seems  to  be  no  means  yet  offered  for  replacing  doors  or  dis- 
pensing with  them  entirely,  there  is  no  question  as  to  their 
objectionable  nature.  They  are  leaky,  and  offer  opportunity 
for  negligent  drivers,  who  by  leaving  them  open  divert  the  current 
from  its  proper  course  or  stop  it  entirely.  The  most  objection- 
able feature  is  their  liability  to  destruction  by  explosion,  and  the 
consequent  annihilation  of  the  current  at  the  most  critical 
time. 

The  Effect  of  an  Obstruction  in  the  Airway. — Any  irregularity 
in  the  sectional  area  of  the  airways  or  any  fall  of  rock  restricts 
the  passage  of  the  air.  So  the  movement  of  the  cages  in  the 
shaft  will  disturb  the  current,  and  the  same  occurs  with-  the  pas- 
sage of  trains.  Their  effect  is  to  increase  the  mine  resistance, 
which  will  be  apparent  at  the  fan  in  an  increased  speed,  even 


464  MANUAL  OF  MINING. 

though  the  quantity  of  air  being  delivered  may  be  less.  If  doors 
are  left  open  in  the  mine,  the  reverse  takes  place. 

When  the  number  of  ventilating  circuits  is  large,  the  obstruc- 
tion in  one  of  the  splits  will  have  little  effect  on  the  fan  or  fur- 
nace as  revealed  by  the  water-gauge.  The  greater  the  number 
of  splits,  the  less  will  be  its  influence.  But  a  redistribution  takes 
place  throughout  the  mine,  and  all  the  other  splits  will  receive 
an  undue  proportion  of  air. 

The  regulator-doors  are  then  reset  until  the  obstruction  has 
been  removed.  The  effect  of  placing  a  regulator  in  any  district 
cannot  be  calculated  without  taking  into  consideration  the  whole 
circumstances  of  the  mine.  Any  change  in  the  position  of  a 
shutter  affects  the  other  airways. 

Precisely  the  same  formulae  will  be  employed  for  the  equiva- 
lent orifice  of  the  mine  as  will  apply  to  mine  regulators,  the 
coefficient  for  the  vena  contracta  being  taken  at  0.65. 


FIG.  1 68. — A  Regulator  Slide-door. 

Safety-doors. — In  some  collieries  an  excellent  device  has  been 
introduced.  This  consists  of  a  sheet-iron  door  hinged  and 
suspended  from  the  roof  at  appropriate  places,  to  be  released 
when  the  explosion  occurs,  drop,  and  close  the  opening,  thus 
replacing  those  disturbed  by  the  accident  and  maintaining  the 
direction  of  the  current  at  this  critical  time. 


DISTRIBUTION   OF   THE  AIR.  465 

Brattices  are  used  as  temporary  expedients  for  subdividing 
any  room  or  airway  in  such  a  manner  as  to  carry  an  inlet  current 
through  the  wider  compartment  to  the  breast  or  face  of  the  work, 
and  return  it  in  the  smaller  section  to  the  return  airway.  Brat- 
tices may  be  made  of  planks  nailed  on  props  at  suitable  distances 
apart,  with  the  interstices  between  the  plank  lathed,  and  the 
whole  tarred  or  calked  with  oakum  to  constitute  an  air-tight 
partition,  or  the  brattices  may  be  made  of  canvas  unrolled  hori- 
zontally and  suspended  from  the  roof  and  frequently  adjusted 
upon  posts,  according  to  the  ventilating  pressure  in  the  airway 
or  room  in  which  they  are  placed.  To  make  canvas  imperme- 
able to  air,  it  is  usually  soaked  with  tar,  though  its  stench  has 
resulted  in  its  substitution  by  an  incombustible  material,  such  as 
asbestos  or  a  soluble  silicate. 

Aid  to  Ventilation. — Other  methods  of  providing  the  return- 
or  inlet-way  consist  in  cutting  a  ditch  in  the  floor  of  a  gallery  and 
boarding  it  over;  in  providing  a  "top  sollar,"  the  timbered  roof 
of  the  gangway  having  a  little  extra  space  for  the  passage  of  the 
air-current  above  it  instead  of  below  it;  in  laying  wooden  air- 
boxes,  metal-pipes,  or  large  canvas  hose  at  one  side  for  tempo- 
rary expedients. 

In  the  rooms,  entries,  and  shafts  the  distance  between  the 
working  place  and  the  last  connection  with  the  airway  must  not 
exceed  22  yards,  and  that  from  the  end  of  the  brattice  must  not 
exceed  one  fifth  of  that  distance. 

"Throughs"  or  "dog-holes"  are  driven  at  every  30  yards 
through  the  pillar  coal  to  connect  rooms,  and  to  allow  of  air 
circulation  through  them  in  turn  as  the  room  advances. 

For  intensifying  the  permanent  air-current  led  to  a  working 
place,  or  for  the  separate  ventilation  of  workings  in  seams  with 
slight  disengagement  of  gas,  compressed  air  may  be  employed 
or  hand-worked  fans  may  be  used  for  separate  ventilation,  which 
is  always  brought  up  so  near  to  the  working  place  that  the  air 
be  not  too  much  diffused. 

EXAMPLE. — A  downcast  6'  X 1 1',  an  upcast  6'  X 12',  300  feet  deep,  supply  air 
to  a  mine  having  a  gangway  3000  feet  long,  50  feet  sectional  area 


4^6  MANUAL  OF  MINING. 

five  splits  receive  each  a  portion  of  the  40,000  cubic  feet  moving.  Required 
the  several  amounts  delivered  to  the  panels  through  the  resistance  of  6'X8', 
700  feet  long;  5'Xf,  looofeet;  s'X8',  1200  feet;  6'X6',  750  feet;  and  4'  X  f, 
800  feet. 

By  substitution  it  will  be  found  that  the  pressure  required  to  overcome  the 
downcast  resistance  is  1.23  Ibs.  per  square  foot.  In  the  mine  the  pressure 
will  be  the  same  throughout;  hence  for  each  split 


which,  solved  for  each  division,  gives  volumes  q=  i6,o59V^;  go^o'^/J;  9679 
V/>;  io,9O9V^T;  and  7543  Vpt  respectively.  The  sum  of  these  equals  the 
total  quantity,  Q,  from  which  we  get  p,  equal  to  0.87  Ib.  Assuming  that  the 
splits  are  made  as  near  as  possible  to  the  downcast  entry  and  reunion  to 
upcast,  which  by  the  way  is  advised,  there  need  be  no  further  allowance  for 
resistances  other  than  those  of  sudden  turns  or  contractions.  The  distribu- 
tion of  the  air.  q,  for  each  district  is  12,070,  6785,  7275,  8200,  and  5670  cubic 
feet. 

The  upcast  offers  a  resistance  to  the  exhaust-air,  the  volume  of  which  is 
greater  than  the  40,000  cubic  feet,  because  of  accretions  from  "blowers," 
moisture,  etc.  Disregarding  these  increments,  the  volume  to  be  exhausted  is 
carried  up  at  a  velocity  of  555  feet  per  minute,  whence  p  is  i.oo  Ib. 

-  The  total  pressure  to  be  imparted  by  increasing  the  downcast  barometric  or 
rarefying  the  upcast  is  therefore  3.01  Ibs.,  requiring  M  of  57  feet,  or  a  dif- 
ference in  temperature  of  about  100°  at  B=  25  inches.  If  the  splits  are  made  at 
stated  distances  along  the  main  gangway,  an  allowance  must  be  made  for  each 
of  the  several  losses  of  friction  in  the  various  lengths  thereof,  remembering 
that  each  branch  split  reduces  the  volume  passing  through  the  remaining  por- 
tion of  the  gallery,  and  correspondingly  the  friction  therein.  It  would  require 
3.7  H.P.  to  do  the  work  upon  the  air;  a  fan  of  43  per  cent  efficiency  would 
necessitate  a  9  H.P.  engine. 

The  calculation  for  the  ventilation  of  a  railroad  tunnel  is  similar.  Assume 
a  tunnel  4961  feet  long;  sectional  area,  336  square  feet.  Then  4961X336= 
1,666,900  cubic  feet  of  air  to  be  changed  every  ten  minutes.  Velocity  of  cur- 
rent, 496  feet. 

p=-  -  =  4.8  Ibs.  per  square  foot. 

A  fan  20'  X  6'  at  forty  revolutions  easily  meets  this  demand.  The  fan  may  be 
applied,  with  or  without  brattice,  at  either  end  of  the  tunnel,  but  this  is  a 
delicate  matter.  It  would  be  better  to  place  the  fan  at  the  mouth  of,  or  a 


DISTRIBUTION   OF   THE  AIR.  467 

furnace  at  the  bottom  of,  one  of  the  connecting  shafts  used,during  construc- 
tion, and  block  the  others  off. 

Required  the  pressure  and  power  to  get  10,000  cubic  feet  of  air  through  air- 
ways aggregating  6300  feet  in  length,  of  which  5000  feet  are  along  a  heading 
5'X8'  in  area,  with  1000  feet  of  airway  6'Xg'  at  the  intake  end  and  300  feet 
of  a  2'X5'  air-course  at  the  upcast  end. 

REFERENCES. 

Mine  Doors,  Coll.  Eng.,  July  1886,  280;  Balance  Doors  and  Air  Hinges, 
Coll.  Eng.,  April  1896,  198;  Stoppings  on  Main  Roads,  E.  Wain,  Fed.  Inst. 
M.  E.,  VI;  Doors,  M.  &  M.  J.,  Vol.  XX,  272. 

Overcasts,  Illustrated,  Mine  Insp.,  Pa.,  1881,  74;  Breakthroughs,  Neces- 
sity for,  Ohio  Mine  Inst.,  1892,  18;  Check  Curtain,  Mine  Insp.,  Ky.,  1891,  61. 

Improvements  in  Ventilation,  M.  &  M.,  Vol.  XXIV,  312;  Air  Column 
in  Ventilation,  M.  &  M.,  Vol.  XX,  165. 

Resistances  of  Air-currents  in  Mines,  Trans.  M.  &  M.  Eng.,  XLV,  62; 
Resistance  of  Air-currents,  T.  L.  Ewen,  Inst.  M.  E.,  X;  The  Effect  of  an 
Obstruction  in  the  Airway,  T.  L.  Ewen,  Inst.  M.  E.,  IX;  Ventilating,  Splitting 
Air-currents,  Door,  and  Valve,  Pa.  Mine  Insp.,  889,  318. 

Saftey-lamp  Glasses,  Coll.  Guard.,  LXXIX,  369;  Experiments  in  Safety 
Lamps,  Coll.  Guard.,  Vol.  LXXIX,  218;  Acetylene  Mining  Lamp,  ColL 
Guard.,  Vol.  LXXXII,  1127;  Detecting  Lamp  Leakage,  Coll.  Guard.,  Vol. 
LXXXIII,  126;  Apparatus  for  Testing  Lamps,  Coll.  Guard.,  Vol.  LXXXVI, 
842;  Tests  of  Safety  Lamps,  Coll.  Guard.,  Vol.  LXXXVI,  261. 


CHAPTER  XV. 

THE  ILLUMINATION  OF  MINES. 

Illumination  in  Mines. — In  an  atmosphere  free  from  explo- 
sive gas,  the  illumination  may  be  had  by  any  form  of  naked 
light.  In  all  metal-mines  candles  are  used,  and,  occasionally, 
the  torch  and  kerosene-lamp.  In  bituminous  mines  known  to 
be  non-gaseous  the  latter  is  employed;  but  in  all  mines  which 
are  at  all  likely  to  develop  gases  the  lamp-flame  must  be  pro- 
tected from  direct  contact  with  the  white  or  fire-damp.  The 
simplicity  of  the  means  of  illuminating  metal-mines  tends  to 
promote  the  comfort  and  safety  of  those  engaged  in  them  as  the 
rays  of  light  from  the  open  candle  or  lamp  are  shed  in  all  direc- 
tions, thus  revealing  dangerous  roofs,  sides,  or  timbering  much 
better  than  any  form  of  safety-lamp  can.  But  only  in  very  shal- 
low collieries  should  this  be  permitted,  and  then  only  after  rigid 
inspection  by  the  fire-boss. 

Candles  and  Oil-lamps. — The  candles  which  are  used  in 
metal-mines  are  usually  of  stearic  acid,  of  which  Proctor  and 
Gamble's  are  the  most  uniform  and  will  best  withstand  the  tem- 
perature of  the  heated  atmosphere.  They  are  cheaper  illu- 
minators than  lamps  in  rooms  and  stopes,  but  not  in  haulage- 
ways.  The  consumption  averages  three  candles  per  man  per 
shift. 

The  candle  may  be  placed  in  a  loop  sewn  in  the  cap  or  the 
pointed  shank  may  be  stuck  into  the  timber,  roadside,  or  floor, 
where  it  is  shielded  from  the  current  of  air.  With  a  proper  sup- 
ply of  pure  air  it  burns  without  unpleasant  smoke  emission,  and 
in  places  where  there  is  a  certain  proportion  of  carbonic  acid  gas 
in  the  air,  which  would  surely  cause  its  extinction  if  left  undis- 

468 


THE  ILLUMINATION  OF  MINES.  469 

turbed,  it  may  be  kept  alight  by  spreading  out  the  wick  so  as  to 
provide  spaces  between  its  cotton  threads,  and  inclining  the 
candle  for  the  tallow  to  feed  the  flame  faster. 

The  common  tin  lamp  with  the  hinged  lid  on  top  and  a  hook 
and  spout  on  either  side — from  the  spout  the  wicking  projects  and 
is  warmed — is  a  more  brilliant  illuminator,  and  is  also  used 
in  coal-mines,  giving  a  moderate  light  of  about  four  candle- 
power,  with,  however,  the  objection  that  it  smokes. 

At  many  collieries  the  pit-bank  and  the  main  roads  from  the 
pit-bottom  are  lighted  by  electricity,  or  by  gas,  manufactured  in 
the  colliery  yard,  but  where  this  is  not  so,  larger  and  more  pow- 
erful lights  than  those  used  at  the  face  are  necessary.  The 
common  arrangement  is  a  can  holding  about  3  to  6  pints  of  oil, 
with  one,  and  often  two  large  wicks,  having  a  tightly  fitting  lid 
and  a  chain  carrying  a  pricker,  for  use  on  the  burning  wick 
when  required. 

Illuminants. — Vegetable  or  animal  lamp-oils  were  at  first 
used  in  safety-lamps,  but  now  mineral  oil  is  commonly  used. 
But  mineral  oils  are  volatile  per  se  without  decomposition. 
Hence,  even  with  heavy  mineral  oils  having  a  flashing-point 
high  enough  to  warrant  their  use  in  safety- lamps,  great  care  is 
necessary  in  their  storage  and  in  the  filling  of  lamps.  Mineral 
oils  are  more  perfectly  fluid,  more  combustible,  and  give  a  better 
light  than  vegetable  or  animal  oils,  but  unless  there  is  a  good 
supply  of  air  to  feed  the  flame  they  are  apt,  owing  to  the  large 
amount  of  carbon  they  contain,  to  deposit  much  soot.  Kerosene 
requires  an  admixture  of  a  less  volatile  oil. 

The  two  staple  illuminants  for  safety-lamps  are  best  refined 
rape-  (colza-oil)  and  seal-oil.  These  give  off  no  vapor  under  a 
comparatively  high  femperature.  The  highest  value  of  these 
oils  is  only  obtained  from  careful  preparation,  and  cheap  produc- 
tions are  to  be  studiously  avoided.  White  lard,  winter-strained 
oil,  is  also  much  used,  the  consumption  being  one  half  gallon  per 
month  for  each  lamp.  In  some  mines  a  mixture  of  equal  parts 
of  seal-oil  and  petroleum  seems  best  to  meet  the  requirements 
of  a  good  illumination  with  a  minimum  of  smoke.  In  a  mine 


47°  MANUAL  OF  MINING. 

using  260  duplex-wick  lamps  the  annual  expense  for  oil,  repairs, 
interest,  etc.,  is  $504.  In  selecting  oil  for  illuminating  pur- 
poses, its  behavior  is  tested  not  only  from  the  standpoint  of  its 
usefulness  as  an  illuminant,  but  also  that  of  its  ability  to  burn 
without  smoke.  When  the  oil  bums  and  the  combustion  is 
perfect,  a  blue  non-luminous  heating- flame  is  produced;  but 
when  the  conditions  are  such  that  the  flame  is  cooled  during 
combustion  or  receives  a  deficiency  of  oxygen,  the  combustion 
is  imperfect,  and  the  portion  of  the  carbon  in  the  oil  is  rendered 
incandescent,  thus  emitting  light.  When  the  oil  is  very  dense, 
the  amount  of  incandescent  carbon  released  becomes  excessive, 
particularly  in  the  presence  of  a  small  amount  of  oxygen,  and 
soot  is  the  result.  The  ideal  oil,  therefore,  should  furnish  a 
maximum  of  light  and  a  minimum  of  soot,  with  sufficient  com- 
bustion to  produce  draught.  A  simple  test  and  a  decisive  one 
may  easily  be  made  for  the  fitness  of  oil  for  use  in  the  miner's 
lamp  by  burning  it,  under  the  ordinary  conditions,  in  a  common 
house  lamp  with  a  short  chimney.  The  mixtures,  which  are 
often  used,  of  mineral  oil  with  animal  and  vegetable  oil  are  always 
objectionable  because  of  the  almost  unendurable  odor,  which 
itself  is  detrimental  to  good  air.  There  is  little  saving  in  their 
employment,  and  they  are  worse  than  the  unadulterated  oil. 
The  very  volatile  oils  and  spirits,  like  benzine,  burn  with  a  clear, 
uniform  flame,  show  an  easily  perceptible  cap  in  the  presence 
of  gas,  and  are  usually  very  sensitive,  being  also  free  from  dan- 
ger in  a  well-constructed  lamp,  even  in  the  hands  of  an  unskilled 
miner.' 

The  Davy  Safety-lamp.  —  Much  ingenuity  has  been  ex- 
pended in  the  endeavor  to  invent  a  safer  means  of  illuminating 
workings  than  that  offered  by  the  naked  flame.  In  1815  Davy 
discovered  that  a  sheet  of  iron-wire  gauze  was  so  good  an  ab- 
sorbent of  heat  that  the  flame  in  contact  with  it  could  not  readily 
pass  through.  Further  experiments  indicated  that  for  mining 
purposes  a  mesh  of  784  holes  to  the  square  inch  was  the  safest, 
and  was  therefore  adopted  as  the  standard.  A  cylinder  of  this 
mesh,  surrounding  the  light,  surmounting  an  oil-lamp  and 


THE  ILLUMINATION    OF  MINES.  471 

capped  by  a  perforated  top,  is  the  form,  which  has  been  little 
changed  since  Davy's  time  (Fig.  169).    After  the  lamp  is  filled 


FIG.  169. — Davy  Lamp  FIG.  170. — Clanny  Lamp. 

with  oil  and  lighted,  it  is  locked,  to  bar  the  miner  against  access 
to  the  flame,  the  wick  of  which  is  trimmed  by  a  wire  passing 
up  through  a  close-fitting  tube  from  the  bottom.  The  combus- 
tion is  supported  by  air  penetrating  the  gauze  at  all  sides. 

Sir  Humphry  Davy  thus  describes  his  invention:  "The  prin- 
ciple of  my  lamp  is  that  the  flame,  by  being  supplied  with  only 
a  limited  quantity  of  air,  should  produce  such  a  quantity  of 
azotic  or  carbonic  acid  gas  as  to  prevent  the  explosion  of  fire- 
damp, and  which,  from  the  nature  of  its  operations,  should  be 
rendered  unable  to  communicate  any  explosion  to  the  outer  air." 
.  Defects  of  the  Davy  Lamp. — This  lamp  has  done  and  con- 
tinues to  do  great  service;  but  it  has  two  defects.  The  first  is 
the  liability  of  the  gauze  to  become  red-hot,  and  allow  the  flame 
to  pass  through  to  the  inflammable  mixture  outside.  The  sec- 
ond objection  is  its  low  illuminating  power.  The  open  spaces 
occupy  only  one  fourth  of  the  area  of  the  gauze,  through  which 
the  light  escapes  horizontally;  still  less  light  gets  out  at  the  top, 
to  illumine  the  roof.  Miners  require  light  thrown  in  every  dircc- 


47 2  MANUAL  OF  111 'N INC. 

tion,  especially  upward.  The  illumination  of  the  roof  is  a  more 
important  matter  than  the  lighting  of  the  working  face.  It  is 
also  more  difficult.  The  Davy  lamp  is  inadequate  for  roof 
inspection.  These  defects  have  been  partially  remedied  in  the 
subsequent  patterns  by  the  use  of  glass,  the  only  impermeable, 
strong,  though  brittle,  transparent  substance. 

The  Clanny  Lamp  (Fig.  170)  is  a  modification  of  the  Davy,  in 
which  a  portion  of  the  wire  cloth  is  replaced  by  a  short  cylinder 
of  glass.  This  improves  the  illuminating  power,  and,  if  enclosed 
in  a  tin  can  or  shield,  becomes  quite  safe  in  a  gaseous  mixture. 
The  majority  of  the  later  forms  of  lamps  use  a  glass  cylinder, 
but  an  internal  explosion,  if  it  occurs,  is  rendered  more  violent  than 
in  the  Davy,  which  offers  less  obstruction  to  the  escaping  gases. 

Stephenson's  Lamp  is  almost  as  popular  in  America  as  are 
the  earlier  forms,  having  a  long  cylinder  of  glass  surrounded  by 
a  wire  gauze,  bonneted  above  by  perforated  copper.  The  air- 
feed  is  also  through  the  gauze,  passing  underneath  and  into  the 
chamber  to  the  flame,  thence  out  at  the  top,  as  usual.  This  plan 
keeps  both  cylinder  and  gauze  cool,  and  its  relative  security 
rests  essentially  on  the  regularity  of  the  draught,  for  if  the  inside 
air  becomes  overheated  the  light  goes  out;  so  it  must  be  sus- 
pended properly.  This  is  an  English  favorite. 

The  Marsaut  Lamp  is  an  improvement  upon  this  form,  and 
stands  a  fair  amount  of  tilting  safely.  With  care,  its  glass  cylin- 
der will  last  three  years  before  breaking.  The  Marsaut  lamp  in 
many  mines  abroad  is  regarded  as 'the  most  suitable  one  for  the 
working  miner,  its  construction  being  simple  and  strong.  As 
an  indicator  of  gas  it  is  reliable,  furnishing  also  a  good  light. 
In  370  in  use,  the  average  consumption  of  rape-seed  oil  was  2 
gallons  per  year.  This  lamp  was  brought  very  prominently 
before  the  public  by  the  Accidents  in  Mines  Commission.  A 
great  difficulty  is  experienced  in  relighting  it,  and,  from  the  wind- 
ing path  pursued  by  the  feed-air,  proper  circulation  does  not 
take  place  until  the  lamp  gets  hot. 

The  Mueseler,  a  Belgian  lamp,  is  like  Dr.  Clanny's,  having 
in  addition  a  conical  chimney  centrally  above  the  flame.  It  is 


THE  ILLUMINATION   OF  MINES.  473 

highly  recommended  in  Europe,  but  must  be  carefully  handled. 
It  does  not  burn  well  in  "dampy"  or  slow  currents.  The  bon- 
neted Mueseler,  an  English  improvement,  is  receiving  the  high- 
est encomium  for  use  in  fiery  mines  and  high  velocity. 

The  Hepplewite-Gray  Lamp  admits  air  at  the  top,  down 
four  tubes,  and  through  an  annular  chamber  above  the  oil  vessel. 
The  only  gauze  employed  is  that  covering  the  outlet  and  annular 
inner  chamber.  A  serious  difficulty  with  it  is  its  liability  to  be 
extinguished  when  suddenly  lowered.  It  undoubtedly  gives 
more  useful  illumination  than  any  other  lamp,  and  as  an  indica- 
tor of  gas  undoubtedly  ranks  superior  to  all  others — except,  pos- 
sibly, the  Pieler  or  Wolf  varieties.  All  other  forms  with  the 
inlet  above  the  glass  will  miss,  say,  4  inches  of  gas  lying  imme- 
diately against  the  roof,  except  when  they  are  tilted  very  much, 
and  then  there  is  great  danger  of  their  going  out.  Many  lamps 
are  now  constructed  to  take  air,  if  desirable,  from  the  top,  like 
the  Gray,  and  thus  also  to  detect  thin  layers  of  gas;  but  even 
then  they  will  not  do  it  so  rapidly.  It  is  possible  to  put  some 
modern  lamps  into  gas  and  take  them  out  again  without  any 
indication  being  given — if  the  test  is  conducted  hurriedly.  This 
is  quite  impossible  with  the  Gray,  as  the  flame  immediately 
'"spires"  up.  Owing  also  to  the  large  amount  of  useful  light 
given  by  it  and  the  way  this  is  directed  on  the  roof,  in  addition  to 
its  delicate  indications  of  gas,  this  lamp  is  preferred  to  all  others 
for  use  by  deputies,  firemen,  timberers,  and  fire-bosses. 

The  Dick  Patent  Port-hole  Lamp  compels  all  the  air  enter- 
ing the  lamp  to  go  immediately  to  the  flame,  thus  losing  no  air, 
and  is  capable  -of  burning  in  a  stagnant  atmosphere.  The  air 
entering  the  lamp  above  the  case  passes  through  the  gauze, 
thence  descends  to  the  flame,  while  the  products  of  combustion 
arise  inside  the  lamp,  to  be  emitted  through  circular  holes  at  the 
top  of  the  bonnet.  The  bonnet  is  made  of  a  seamless  steel  tube, 
and  is  light  and  strong. 

The  Clifford  Lamp  is  a  new  one  and  has  many  excellent 
points.  The  light  given  is  good.  A  plentiful  supply  of  air 
enters  the  lamp.  It  illuminates  the  roof  as  well  as  the  sides  of 


474  MANUAL  OF  MINING. 

the  workings.  It  is  not  sensitive  to  tilting  or  violent  movement. 
The  bonnet  arrangements  prevent  the  possibility  of  a  current  of 
any  velocity  from  impinging  directly  upon  the  gauze,  and  thereby 
render  the  lamp  safe  in  explosive  currents. 

The  Beard-Mackie  Lamp  has  a  small  brass  disc  supporting  a 
H-bent  rod,  across  whose  standards  are  strung  platinum  wire. 
The  height  of  the  flame  is  made  visible  by  rendering  the  wire 
incandescent.  The  percentage  of  gas  present  is  known  from 
the  height  of  wire  rendered  luminous.  As  the  wires  can  be 
altered  in  position  by  the  mine-boss,  the  warning  line  can  be 
regulated  according  to  the  risk  to  be  run.  The  lamp  is  enclosed 
in  a  wire  gauze  or  in  a  bonneted  glass  cylinder. 

The  Woolf  Benzine  Safety-lamp  is  an  emphatic  departure 
from  the  varieties  above  described,  in  that,  first,  it  burns  benzine 
or  naphtha;  second,  it  contains  a  patent  self- igniter  capable  of 
relighting  the  lamp  fifty  times  without  opening;  and,  third,  it 
contains  a  locking  device  which  it  is  impossible  to  open  except 
by  the  use  of  an  exceedingly  powerful  magnet.  This  lamp, 
because  of  the  sensitiveness  of  its  illumination,  is  a  delicate  de- 
tector of  gas,  and  has  met  with  very  ready  acceptance  throughout 
coal-mining  districts,  there  being  possibly  80,000  in  use  in  Ger- 
many. It  furnishes  a  good  indication  of  the  presence  of  gas 
by  the  height  of  its  flame.  A  percentage  of  light  carburetted 
hydrogen,  varying  from  i  to  5  per  cent  in  the  total  mixture,  will 
increase  the  height  of  the  flame  from  2.6  to  5.8  cm.  If  the 
flame  be  partially  turned  down,  a  still  greater  degree  of  sensi- 
tiveness will  be  manifested. 

The  Requisites  of  a  Safe  Lamp. — The  lamp  must  be  self- 
contained,  strong,  portable,  and  not  heavy,  require  little  atten- 
tion from  the  miner  during  twelve  hours  of  sustained  light,  and 
capable  of  being  placed  in  any  position,  besides  giving  perfect  insu- 
lation from  the  fiery  gas.  The  use  of  the  safety-lamp  is  to  secure 
protection  to  the  miners  from  explosive  mixtures,  whether  by 
allowing  of  safe  examinations  by  the  fireman  before  the  shift 
begins,  or  by  preventing  explosion  from  sudden  eruptions  or 
ordinary  discharges  of  gas  during  the  shift.  It  should  afford 


THE  ILLUMINATION  OF  MINES.  475 

good  illumination  of  the  roof  during  repairs  to  the  latter,  or  dur- 
ing ordinary  operations  of  mining,  without  tempting  the  miner 
to  uncover  the  flame.  The  wire-mesh  casing  reduces  the  pho- 
tometric value  of  the  light  to  a  low  degree. 

The  illumination  from  any  of  these  lamps  is  very  feeble.  It 
is  less  in  any  direction  than  the  horizontal.  Of  all  the  lamps 
the  Gray  sends  the  best  light  upward.  The  candle-power,  hori- 
zontally, of  the  Roberts  is  highest — about  18,  and  of  the  Clanny 
the  lowest — nearly  6.  On  this  account  a  lamp  must  be  able  to 
be  held  tilted  without  extinguishment,  and  be  unaffected  by 
violent  oscillations.  The  conditions  dictated  by  safety  circum- 
scribe the  lines  of  attempted  improvement  in  the  degree  of  illu- 
mination. For,  with  a  large  mesh,  the  lamp  is  incapable  of  pre- 
venting high-speed  gaseous  currents  from  entering  the  con- 
struction chamber  and  blowing  the  flame  against  the  opposite 
gauze,  or  forcing  gas  into  contact  with  the  flame.  The  safe 
lamp  must  have  a  mesh  impermeable  to  the  high-speed  currents. 
Glass  improves  the  illumination  but  introduces  other  dangers — 
risk  of  breaking,  if  unannealed,  by  cold-air  current  or  dripping 
water,  and  the  difficulty  of  securing  a  good  joint  to  the  gauze 
cylinder. 

Of  the  forty-one  explosions  which  occurred  in  a  certain  dis- 
trict during  1896,  in  four  cases  the  immediate  cause  of  ignition 
was  referred  to  a  naked  light  or  to  a  deterioration  of  the  safety- 
lamp;  in  twenty-five,  to  the  passing  of  the  safety-lamp  flame 
against  the  gauze  through  some  careless  movement,  too  high  a 
speed,  or  "falls  in."  The  remaining  twelve  were  from  shot; 
fire,  or  other  undetermined  causes. 

Designers  of  lamps  have  difficulties  to  overcome  other  than 
these.  They  include  the  rough  usage  lamps  receive  in  the  mine, 
the  violent  shaking  and  tilting  they  are  subjected  to,  frequently 
causing  discharges  from  the  oil-chambers;  the  dust  finely  sus- 
pended in  the  air,  or  settling  on  them  at  the  faces,  making  them 
more  explosive;  oblique  explosive  currents  at  the  face  caused  by 
falling  stone  or  coal,  close  to  the  lamps;  and  the  grimy  hands  of 


476  MANUAL  OF  MINING. 

the  collier,  which  clog  the  gauzes  and  inlet  feed-holes  with  dirt 
and  grease. 

An  absolutely  safe  lamp  is  therefore  difficult  to  obtain,  and 
notwithstanding  the  various  modifications,  there  is  as  yet  none. 
The  lamp  which  cannot  ignite  in  an  explosive  mixture  outside 
of  it  is  yet  to  be  invented. 

The  defects  of  the  Marsaut  lamp  are  at  the  metallic  connec- 
tion of  the  bonnet  with  the  other  parts.  They  heat  up  from  the 
impingement  of  the  flame  and  gases.  Gas  moving  at  50  feet  per 
second  can  penetrate  the  two-gauze  Marsaut,  but  a  three-gauze 
pattern  will  resist  the  current.  The  only  source  of  danger  in 
the  Mueseler  lamp,  not  common  to  the  other  types,  is  its  strong 
tendency  to  smoke;  otherwise  it  is  a  most  efficient  lamp,  show- 
ing a  bright  and  steady  flame  in  the  strongest  current.  The 
Gray  lamp  presents  the  risk  of  the  gas  burning  at  the  cylindrical 
strip  of  gauze  under  the  glass,  which  in  low-speed  currents  heats 
the  lower  edge  of  the  glass  strongly,  and  in  high-speed  currents 
allows  the  heated  gases  to  pass  across  against  the  glass  on  the 
other  side.  The  top  of  the  lamp  can  be  easily  tampered  with, 
and  the  gauze  at  the  outlet  is  liable  to  be  obstructed  by  soot  if 
the  flame  should  smoke. 

It  is  worthy  of  note  that  the  brass  lamps  are  less  bright  than 
the  iron  lamps  of  the  same  pattern.  A  round  wick  is  not  so 
luminous  as  a  broad,  flat  wick  for  a  given  oil  consumption.  Seal- 
oil  is  photometrically  better  than  rape-seed-oil  illuminant.  The 
insufficient  light  of  a  safety-lamp,  combined  with  the  difficult 
and  trying  conditions  of  the  bonneted  forms,  is  proving  injuri- 
ous to  the  eyesight  of  miners,  which  serious  evil  is  growing. 
Photophobia  is  rare  where  candles  are  used,  or  where  the  lamp 
is  hung  behind  the  miner. 

Bonneted  Lamps. — In  order  to  be  safe  in  the  highest  velocity 
of  air-currents  within  a  given  mine,  the  flame  must  be  enclosed 
not  only  in  a  wire  gauze,  but  also  in  a  more  or  less  impermeable 
hood  or  bonnet,  while  the  inlet  area  for  the  feed-air  must  be 
reduced  to  the  smallest  allowable  dimensions.  Many  lamps 
now  exist  which  appear  to  resist,  in  a  highly  explosive  atmos- 


THE  ILLUMINATION    OF  MINES.  477 

phere,  current  velocities  up  to  3000  feet  per  minute  for  a  period 
of  several  minutes;  and  the  four  lamps  which  were  brought  to 
the  attention  of  the  Mines  Accident  Commission,  which  re- 
ceived special  attention  for  their  security,  illuminating  power,  and 
simplicity  of  construction,  were  the  H.-Gray,  Marsaut,  bonneted 
Mueseler,  and  Thomas's  modification  of  the  bonneted  Clanny. 

The  bonnet  screens  the  gauze  cylinder  from  the  effects  of 
draughts  that  blow  the  flame  through  the  meshes  and  set  up 
a  fiery  heat  by  the  excessive  air  and  gas  that  enter  above  the 
flame  of  the  wick.  It  limits  the  supply  of  air  to  that  required 
for  the  oil-flame  only.  Such  bonneted  lamps,  whose  flames  are 
protected  from  the  direct  effects  of  the  strong  ventilating  current, 
may  be  used  with  safety  for  illumination  in  mines  producing 
fire-damp.  Even  in  dry,  dusty  mines  also  developing  fire-damp 
some  of  these  lamps  are  safe,  though  not  all,  for  many  well- 
authenticated  cases  of  failure  are  recorded  where  the  dust  has 
proven  fine  enough  to  pass  through  the  gauze  meshes,  to  be 
reduced  to  the  state  of  incandescence  in  the  inner  chamber. 
Locked  safety-lamps  arc  insisted  upon  in  mines  when  cutting 
through  clay- veins  in  solid  workings.  The  Hepplewite-Gray 
and  the  bonneted  Mueseler  have  the  best  resistance  to  explosive 
currents  of  high  velocity,  and  the  South  Side  Committee  report 
the  following  relative  speeds  at  which  the  respective  lamps  and 
the  air-current  can  safely  pass:  Davy,  360  feet  per  minute; 
Clanny,  600  feet;  Stephenson,  780;  Mueseler,  naked,  1200; 
Mueseler,  bonneted,  2400;  Marsaut,  in  a  can,  2440;  and  the 
Davy,  in  a  shield,  2400.  The  North  of  England  Institute  of 
Mining  Engineers  gives  the  safe  velocities  at  720,540,  and  the 
others  higher.  The  British  Royal  Commissioners  of  Accidents 
approved  the  Gray,  Marsaut,  and  the  bonneted  varieties  as  safe  at 
high  speeds.  The  'common  Davy  or  Geordie  lamps  are  unreliable. 

Locking  Lamps. — All  safety-lamps  have  locks  to  them.  The 
oldest  form  of  lock  is  still  used  on  most  lamps,  notwithstanding 
its  acknowledged  insecurity.  It  consists  of  a  screw-bolt  with 
a  square  head  which  is  turned  by  a  key  until  it  has  entered  a 
hole  bored  in  the  oil-chamber,  which  is  supposed  to  prevent 


478  MANUAL  OF  MINING. 

it  from  being  unscrewed  until  the  screw-bolt  is  withdrawn.  But 
this  may  be  done  by  a  couple  of  nails  filed  to  fit  the  bolt,  by 
pieces  of  wire  bent  for  the  purpose,  or  even  by  the  point  of  a 
pick.  It  affords  a  little  more  security  than  the  original  device  of 
the  screw-lock. 

The  importance  of  locking  the  lamp  so  that  its  flame  can- 
not be  exposed  to  the  gas  is  manifest,  as  there  are  many  temp- 
tations to  the  miner  to  open  it.  Most  of  them  can  be,  and 
are,  opened  by  easily  extemporized  mechanical  means,  while 
others  are  rather  more  difficult  to  open. 

•All  manner  of  permutation-locks  and  magnetized  plates  are 
offered  on  the  market,  besides  the  lead-plug  seal  with  which 
the  lamp  is  riveted  after  each  filling.  On  account  of  its  sim- 
plicity and  ease  of  treatment,  as  well  as  from  the  measure  of 
security  it  affords,  the  lead-pin  is  coming  more  and  more  into  use 
at  collieries.  The  pin  is  moulded  with  a  head  at  one  end,  and 
fits  openings  in  the  two  parts  of  the  lamp  to  be  held  together. 
When  in  place  it  is  firmly  riveted  and  punched  with  some  device 
at  both  ends.  This  forms  the  means  of  detection,  if  the  lamp 
should  be  wrongfully  opened. 

Magnetic  locks  are  employed  to  advantage.  Lamps  so 
fitted  cannot  be  opened  except  with  the  aid  of  a  powerful  mag- 
net. They  resist  all  the  efforts  of  the  miner,  short  of  wreckage 
of  the  lamps.  The  Woolf  lamp  and  the  Craig  &  Bidder's 
lamps  are  so  fastened. 

Extinguishing-locks.  These  are  devised  to  remove  all  temp- 
tation from  the  collier  to  unlock  his  lamp  by  so  contriving  it 
that  the  process  of  unlocking  also  extinguishes  the  flame. 

Lighting  and  Relighting  Locked  Lamps.  —  A  great  many 
lamps  become  extinguished  at  the  working  faces  from  a  variety 
of  causes,  and  this  has  led  to  designs  of  lamps  which  may  be 
quickly  relighted  without  their  being  unlocked.  The  designs 
may  be  divided  into  two  kinds,  in  one  of  which  lamps  may  be 
relighted  anywhere  by  the  user;  the  other,  only  by  an  authorized 
person  at  properly  appointed  stations  by  the  application  of 
electricity. 


THE  ILLUMINATION   OF  MINES.  479 

Woolf  s  lamp  belongs  to  the  first  class,  the  relighting  arrange- 
ment being  adapted  for  the  volatile  illuminant,  benzoline,  used 
in  it.  The  lamp  contains  a  reel  of  paper  on  which  fulminating 
spots  are  placed  at  intervals  of  about  a  quarter  of  an  inch,  er.ch  of 
which  can,  in  turn,  be  brought  opposite  the  wick,  and  at  the 
same  time  be  struck  by  a  small  spring-hammer,  operated  by  a 
button  at  the  bottom  of  the  lamp;  the  composition  explodes 
and  ignites  the  benzoline  vapor  surrounding  the  wick. 

Use  of  Lamps. — Safety-lamps  provided  with  the  best  form 
of  lock  to  prevent  their  being  tampered  with  by  ignorant  or 
reckless  workmen,  and  thoroughly  tested  before  given  out,  will, 
if  carefully  used,  afford  considerable  protection  in  the  mine.  No 
lamp  is  safe  unless  kept  in  thorough  repair,  and  any  infraction 
of  rules  regarding  careful  use  should  be  severely  punished. 

Examining  and  Testing  Lamps. — The  lamp-room  is  on  the 
surface,  where  it  is  safer  than  underground.  The  safety-lamps 
are  complicated  in  their  construction,  and  the  work  of  examin- 
ing and  repairing  them  in  large  collieries  is  considerable.  As 
it  is  difficult  for  the  eye  to  detect  defects  in  adjustment,  there 
should  be  provided  some  simple  means  for  reliably  testing  the 
joints  between  the  glass  cylinder  and  the  gauze.  It  is  well  to 
test  each  lamp,  before  it  is  taken  into  the  pit,  by  placing  it,  lighted, 
in  an  explosive  mixture. 

Cleaning  Lamps. — At  small  collieries  having  a  few  safety- 
lamps  in  use  the  cleaning  is  usually  done  •  by  hand,  without 
the  aid  of  mechanical  contrivances  to  unscrew  or  remove  the 
internal  fittings;  an  ordinary  hand-lamp  brush  is  used  to  rub 
the  dirt  from  the  gauzes.  Even  when  the  workmen  clean  their 
own  lamps  there  is  at  least  one  lamp-keeper  at  each  colliery 
to  replenish  the  oil  vessels,  renew  the  wicks,  and  replace  worn 
washers  or  broken  glasses,  except  in  the  very  rare  instances  in 
which  the  men  themselves  supply  all  safety-lamps  and  after- 
wards maintain  them  in  a  perfect  state  of  repair.  Where  oil 
is  burned  the  gauze  should  be  steeped  in  a  hot  alkaline  solution, 
to  free  it  of  soot,  etc.  Lamps  burning  benzine  are  not  clogged 
with  carbonaceous  deposit. 


480  MANUAL  OF  MINING. 

To  avoid  waste,  manufacturers  furnish  automatic  fillers 
holding  the  exact  quantity  sufficient  for  one  lamp. 

Electric  Lights  in  Mines. — The  advantages  of  electricity 
for  lighting  about  a  mine  are  its  decreased  cost,  better  illumina- 
tion, absolute  reliability,  and  greater  freedom  from  accident. 
The  lights  outside  are  arc  lamps;  those  inside  are  incandescent. 
Arc  lamps  are  either  of  the  open  type,  or  enclosed.  The  latter 
give  less  light  under  similar  conditions,  but  their  distribution  is 
more  even.  The  former  require  from  8  to  10  amperes,  and  the 
latter  5  amperes  of  current.  Though  the  efficiency  of  the  latter 
is  less,  their  consumption  of  carbon  and  the  expense  of  their 
renewal  are  also  less. 

The  lamps  may  be  connected  continuously  in  one  series,  in 
which  case  the  dynamo  is  series-wound.  If  the  lamps  are  in 
series,  each  one  is  provided  with  an  automatic  cut-out  allowing 
the  others  to  receive  the  current  in  case  of  its  failure.  When 
connected  in  several  short  circuits  from  the  one  main,  "in  paral- 
lel," they  will  require  a  compound- wound  machine.  A  resist- 
ance is  then  put  into  series  with  each  circuit  in  order  to  keep  its 
current  at  the  proper  pressure.  The  latter  connection  is  more 
common  for  mining  work. 

Each  circuit  is  provided  with  a  fuse  which  bears  the  same 
relation  to  the  installation  that  the  safety-valve  does  to  the  boiler* 
Fuses  are  made  of  alloys  of  tin  or  lead  for  the  5o-ampere  circuits, 
and  of  copper  for  heavier  lines. 

The  light  wires  are  protected  from  damp  by  being  inserted 
in  cast-iron  piping  or  lead  tubing.  A  double-pole  switch  mounted 
conveniently  on  a  metal  base,  easily  reached,  opens  and  closes 
the  circuit  for  the  lamps.  Having  two  contacts,  it  divides  the 
current  and  the  spark,  and  thus  reduces  the  risk. 

Lamps  of  16  candle-power  on  a  circuit  of  100  volts  require 
0.6  ampere.  On  the  2 20- volt  circuit;  0.27  ampere.  These 
correspond  to  3.75  watts  per  candle-power. 

12  lamps  of  16  C.P.  require  i     H.P. 

6      "      "  32  C.P.       "       i     H.P. 

2  arc  lamps  of    600  C.P.       "       i     H.P. 
2    "       "      "    icoo  C.P.       "       1.7  H.P. 


THE  ILLUMINATION   OF  MINES.  481 

It  is  usual  to  allow  30  candle-power  for  every  100  sq.  ft.  of 
floor  area,  when  they  are  placed  to  advantage.  A  radius  of  70 
feet  is  allowed  for  each  5-ampere  enclosed  lamp  or  a  lo-ampere 
open  lamp. 

For  parallel  distribution  of  current  to  lamps,  let  w=  watts 
per  candle-power;  W=  watts  per  lamp;  £=  voltage  at  the 

W 
lamp   terminals.    Then   /  =  p-=the    current    in    amperes    per 

lamp.  All  wiring  is  based  on  an  allowable  drop,  of  which  5  per 
cent  may  be  considered  a  good  average  for  the  voltage  loss  in 
this  character  of  wiring.  The  area  of  a  conductor  to  furnish 
the  current  for  a  given  number  of  lamps,  N,  is 

10.81  X2DXI     10.81  X2DXN 
A.= 


drop  in  volts      R  Xper  cent  loss* 

R  is  the  heat  resistance  of  one  lamp. 

EXAMPLE. — It  is  desired  to  illuminate  the  grounds  and  buildings  about 
the  tipple  by  100  i6-candle-power  lamps  on  a  line  aggregating  600  feet 
from  the  generator,  30  lamps  of  32  candle-power,  each  at  a  distance  of  750 
feet,  and  6  arc  lamps  in  pairs  at  a  distance  of  600  feet  from  the  power. 
Required  the  electric  horse-power  and  the  size  of  conductor,  the  voltage  of 
the  incandescent  lamps  to  be  200  and  of  the  arc  lamps,  ico. 

As  each  lamp  requires  0.3  ampere  for  16  candle-power,  0.6  ampere 
for  32  candle-power,  and  5  amperes  for  the  arc  lamp,  the  currents  will  be  30, 
18,  and  30  amperes,  respectively.  For  the  small  lamps  600  feet  of  line  must 
carry  30  amperes.  This  will  require  the  wire  of  No.  4  B.  &  S.  gauge  having 
a  resistance  of  1.5  ohms  per  mile.  The  loss  on  this  line  will  therefore  be 
45  volts  per  mile  and  5  volts  for  the  line. 

The  larger  incandescent  lamps  will  be  on  a  cabb  of  No.  6  B.  &  S.  gauge 
having  a  resistance  of  2.5  ohms  per  mile  or  loss  of  43.5  volts  per  mile;  and 
6.2  volts  for  750  feet. 

The  dynamo  should  deliver  a  current  at  its  terminals  of  2.06  volts  pressure 
and  78  amperes.  This  equals  16,068  watts,  which  at  85  per  cent  efficiency 
correspond  to  26-brake  horse-power. 


4^2  MANUAL  OF  MINING. 


REFERENCES. 

Location  of  Regulators,  M.  &  M.,  Vol.  XXII,  501;  Mine  Doors,  Geo. 
Davidson,  111.  Mine  Insp.,  II,  177;  Balanced  Doors  and  Iron  Air  Bridges  as 
Preventatives  of  Mine  Explosions,  W.  M.  Morris,  Coll.  Eng.,  XVI,  198;  Stop- 
pings on  Underground  Roads,  E.  B.  Wain,  Fed.  Inst.  M.  E.,  VI  and  VII; 
Fire  Doors  for  Mine  Shafts,  R.  G.  Brown,  E.  &  M.  Jour.,  Vol.  LVII,  321; 
Illuminating  Oils  in  Mines,  Ohio  Mine  Insp.  Report,  R.  Haseltine,  1895,  46; 
Lamp  Oils,  Am.  Mfr.,  LVI,  918. 

Safety  Lamps,  Comments  by  Mine  Insp.,  Coll.  Mgr.,  1894,  24;  Lamps: 
Discussions,  Remarks  on  Use  of,  Coll.  Guard.,  Dec.  n,  1896,  1117;  Safety 
Lamp,  Development  of,  M.  &  M.,  Vol.  XX,  39;  Safety-lamp  Flames,  M.  & 
M.,  Vol.  XX,  417;  Wolff  Lamp,  M.  &  M.,  Vol.  XXIII,  226;  Eugene  B. 
Wilson,  Amer.  Inst.  M.  E.,  XIII,  129;  E.  J.  Schmitz,  Amer.  Inst.  M.  E.,  XVI, 
410;  Overcasts,  Robert  Mauchline,  Pa.  Rep.  of  Min.  Insp.,  1881,  74;  1885, 
Packet. 

Electric  Lighting  for  Mines,  a  Paper,  Coll.  Mgr.,  1894,  85;  How  to  Light 
a  Colliery  with  Electricity,  S.  F.  Walker,  Brit.  Soc.  Min.  Stud.,  XIII,  147; 
Electric  Lamps  in  Coal  Mines,  E.  &  M.  Jour.,  LIX,  316;  Electric  Lamps, 
M.  &  M.,  Vol.  XXII,  195. 

Ventilation  of  Thin  Workings,  M.  &  M.,  Vol.  XXII,  141;  Ventilation  of 
Deep  Workings,  M.  &  M.,  Vol.  XXII,  273. 


CHAPTER  XVI. 

ACCIDENTS  IN  MINES. 

The  Inherent  Dangers  in  the  occupation  of  mining  are  great. 
The  uncertainties  in  the  overlying  ground,  the  ever-lurking  foe 
in  the  form  of  gas,  the  as  yet  imperfect  methods  of  illumination, 
demand  the  greatest  vigilance  as  the  price  of  comparative  safety. 
Add  to  these  the  ever-present  factor  of  human  carelessness,  and 
the  equation  becomes  one,  containing  two  independent  variables, 
the  solution  of  which  is  hard  to  find.  To  what  extent  the  inher- 
ent dangers  could  be  overcome,  were  the  factor  of  carelessness 
eliminated,  it  is  difficult  to  say.  The  German  Government,  out 
of  a  total  of  7933  lives  lost  in  coal-mining,  attributes  5179,  or 
65  per  cent,  to  the  purely  inherent  dangers  of  the  occupation; 
2397,  or  30  per  cent,  to  the  carelessness  of  the  victims;  248  to 
the  fault  of  fellow  workmen,  and  69,  or  0.9  per  cent,  to  defects 
in  working.  Reports  of  mine  inspectors  attribute  60  per  cent  of 
the  deaths  in  mines  of  the  United  States  to  carelessness;  the 
40  per  cent  to  unavoidable  or  unforeseen  causes. 

Comparative  Hazard  in  Nations. — Under  this  head  it  is 
impossible  to  obtain  absolutely  accurate  conclusions,  owing  to 
the  great  difference  in  methods  in  vogue  in  different  countries. 
In  Europe  the  record  of  accidents,  injuries,  and  fatalities  has 
been  much  more  elaborate  and  precise  than  in  America,  while 
the  classification  of  accidents  varies  in  the  different  countries. 
In  France  an  accident  is  not  entered  on  the  records  unless  the 
person  is  prevented  from  working  for  a  period  of  twenty  days; 
in  Austria  it  is  the  same;  in  Belgium  eight  days  is  the  limit;  in 
Germany  the  law  is  still  more  strict,  as  it  is  necessary  to  give 
notice  of  inability  to  work  for  any  period  exceeding  three  days. 

483 


484 


MANUAL  OF  MINING. 


Every  detail  accompanies  the  inspectors'  reports  in  Europe, 
such  as  the  age  of  victim,  nationality,  character  of  work,  length 
of  time  in  service,  exact  location  of  accident  in  mine,  nature  of 
injury,  character  and  extent  of  medical  attendance,  length  of 
time  incapacitated,  number  of  days  absent  from  work  due  to 
injury.  Even  family  relations  are  reported,  as  number  depend- 
ent upon  the  man,  etc.  In  America  inspectors'  reports  are  as 
a  rule  more  general  or  indefinite,  due  to  the  different  conditions 
and  consequent  differing  laws  existing  in  the  several  States.  As 
yet  we  have  no  Federal  requirements  save  those  locally  applied 
in  territories  and  reservations. 

There  are  two  methods  of  comparison  of  hazards:  ist,  classi- 
fication by  rate  of  accidents  per  number  of  tons  of  coal  produced 
annually;  2d,  by  rate  per  accident  per  number  of  employees 
working  underground,  or  by  total  of  surface  and  underground 
laborers.  If  the  estimate  be  made  on  the  first  basis,  allowance 
must  be  made  for  the  greatest  efficiency  of  the  American  laborer, 
due  largely  to  the  shallowness  of  our  coal-seams  and  the  exten- 
sive use  of  machine  cutters.  In  the  table  below  can  be  seen  the 
productiveness  of  the  American  compared  with  the  English  and 
the  Belgian  miner. 

DEATH-RATES  AND  TONNAGE  PER  EMPLOYEE. 


Tonnage 
per  Em- 
ployee. 

Tons  of 
Coal  per 
Death. 

Rate 

per  1000 
Em- 
ployees. 

United  States,  1893  t< 
«          «         ft     t 

Ohio,                  "     " 

Austria,             1894  ' 
Belgium,           1896  ' 
France, 
Great  Britain,    "     ' 
Prussia,               '  '     ' 

>  1902,  b 

1903 
1902 

Dth  incl 
<          < 

< 

« 

usive,  bituminous 
anthracite. 

660 
381 

5I1 

326 

178 
207 
314 
338 

221,091 
126,795 
292,596 
199,526 
"3,358 
182,525 
238,04? 
150,000 

..89 
5.81 
.76 
.718 

.01 

•23 
•31 
.16 



The  second  method  of  classification  is  generally  accepted  as 
the  more  reliable  and  is  more  readily  understood.  It  also 
affords  a  comparable  standard  of  comparison  with  other 
trades.  The  following  table  requires  little  explanation.  It 
represents  conditions  in  the  three  prominent  coal-producing 


ACCIDENTS  IN  MINES. 


485 


States  of  America.  The  utmost  care  has  been  taken  to  insure 
the  accuracy  of  all  data  used,  and  all  statistics  have  been  offi- 
cially verified,  as  a  thorough  and  comprehensive  knowledge  of 
the  dangers  and  fatal  consequences  in  this  occupation  is  a  matter 
of  absolute  necessity  to  those  engaged  in  it. 


Penns> 

Ivania. 

Anthracite. 

Bituminous. 

Tons  of  coal  
Number  of  employees  inside.  . 
Number  of  employees  outside. 
Fatalities,  total  
Serious  injuries  

36,9  1  1,^4 

93,377 
49,762 
300 
641 

98,946,204 
110,015 
25.371 
*45° 
861 

30,021,300 

4i,5i8 

44,487 

99 
406 

23,929,267 

27,613 
9808 
81 
298 

Deaths  per  1000  employees.  . 
Injuries  per  1000  employees. 
Number  of  days  worked  .... 
Tons  of  coal  per  life  lost  
Annual  output  per  employee.  . 

2.25 
4-32 

220 
123,039 
376 

3-36 
6.36 
116 
217,000 
898. 

2.  I 

8.8 
210.4 

303,245 
724 

2.1 

7-9 
176.4 
293,818 
867 

*  Includes  an  explosion  involving  112  lives. 

The  Death-rate  in  Mines.  —  During  the  earlier  period  of 
mining  the  loss  of  life  annually  was  at  a  rate  of  i  death  for  every 
200  persons  employed.  With  time  this  rate  is  appreciably 
reduced.  In  Great  Britain,  during  1903,  it  was  i  for  every  688 
employees.  The  average  in  the  United  States  is  now  about 
2.3  lives  for  each  1000  men  employed,  or  i  for  every  435.  During 
the  year  1902  the  total  mineral  output  of  the  United  States  re- 
quired nearly  1,000,000  employees,  of  whom  at  least  2500  lost  their 
lives.  In  Great  Britain,  France,  Belgium,  and  Prussia,  during 
the  same  year,  coal-mining,  in  producing  478,459,088  tons  of  coal, 
killed  1937  below  and  306  at  the  surface.  Metal- mines  are  more 
hazardous  than  coal-mines,  but  in  either  the  risk  may  be  said 
to  be  less  than  in  other  of  the  dangerous  occupations.  The  annual 
rate  of  death  among  railway  brakemen  during  the  past  decade 
was  12.5  deaths  per  thousand;  that  among  trainmen,  9.8  per 
1000.  The  highest  rate  in  any  coal  district  to  date  is  8.5  per 
1000 ;  the  anthracite  coal-miners  is  5.6,  and  among  the  anthracite 
laborers  4.6  per  1000.  In  the  United  States  and  Canada,  for 
the  decade  ending  in  1902,  there  were  -1,150  deaths,  a  rate  of  2.Cj 


486  MANUAL  OF  MINING. 

for  each  1000  employees.  In  the  entire  railroad  service  during 
the  same  time  the  average  rate  of  death  was  2.64  for  each  1000 
employees.  Among  the  bituminous  coal-miners  the  rate  aver- 
aged 2.2.  The  accident  liability  inside  the  mine  is,  of  course, 
greater  than  among  the  laborers  at  the  surface. 

Statutory  Provisions  Properly  Governing  Underground  Opera- 
tions, to  provide  safety  to  miners,  require:  A  safe  ingress  and 
exit,  with  ample  means  of  communication  between  the  bottom 
and  the  top;  the  examination  of  the  working  places  by  a  fore- 
man prior  to  the  entry  of  the  men;  copious  volumes  of  air  in 
well-directed  currents  to  dilute  the  accumulations  of  gas;  secure 
timbering  and  effective  illumination  of  the  working  places;  the 
posting  of  copies  of  rules,  advice,  and  precautions,  printed  in 
several  languages;  official  inspection ;  an  ample  supply  of  timbers ; 
large  pillars  to  prevent  caves  of  roof;  supervision  in  the  use  of 
explosives  in  each  working  place;  boundary  pillars  between 
gaseous  sections  of  a  mine  and  drowned  or  abandoned  mines. 
Faithfully  executed,  these  provisions  are  ample  to  protect  the 
men  from  the  extreme  underground  dangers. 

It  cannot  be  denied  that  the  American  death-rate  is  higher 
than  in  European  countries,  nor  does  it  show  the  strong  down- 
ward tendency  that  is  apparent  in  their  statistics.  A  fair  ex- 
ample may  be  the  record  of  Great  Britain  for  the  semi-decade 
since  1857.  The  annual  death-list  showed  nearly  icoo  lives 
sacrificed  in  1852,  and  it  is  a  striking  fact  that  not  many  more 
were  lost  in  1902,  but  the  number  of  employees  have  increased 
during  that  time  from  about  200,000  to  850,000.  The  improve- 
ment in  the  conditions  of  English  mines  is  also  manifest  by  the 
following  summary,  showing  the  tons  of  coal  produced  for  each 
life  lost: 

1852 59,8oo  tons 

1862 82,500    " 

1872 112,000    " 

1882 145,000    " 

1892 200,000       " 

1902 227,280      " 


ACCIDENTS  IN  MINES. 


487 


It  is  true  the  statistics  of  the  States  include  only  the  coal- 
mines, though  it  is  improbable  that  the  inclusion  of  the  metal- 
mines  would  bring  the  rate  even  to  that  of  Great  Britain,  particu- 
larly as  there  are  many  small  coal-mines  not  subjected  to  official 
inspection  whose  records  are  never  kept.  America  is  behind 
the  other  nations  in  the  efficiency  of  its  safeguards,  or,  at  least, 
in  the  results. 

AVERAGE  NUMBER  OF  ACCIDENTS  IN  MINES  IN  THE  UNITED  KINGDOM  IN 
SEMI-DECADES. 


Number  of 
Employees. 

Total 
Deaths. 

Annual  Death-rate  per  1000  Employees. 

Periods. 

1 

•d 

M 

' 

5fii 

it 

II 

Below. 

Above. 

gf 

i 

I 

"o  3 

1 

"  "  ^"3 

^   0 

~S 

1 

0 

a 

£r*> 

<fi 

is 

4%f* 

-M  d 

W 

& 

a 

i 

HU 

H 

H 

1851-55 

182,427 

47,047 

937 

47 

.280 

.016 

i.  206 

0-556 

5.149 

.012 

1-301 

1856-60 

208,763 

53,832 

964 

53 

•234 

.846 

O.SQQ 

0.648 

4.628 

0.994 

3.883 

i86i-6s 

237,779 

8Q8 

68 

.618 

.714 

0.668 

0.790 

3-791 

.105 

3.240 

1866-70 

269,813 

69,574 

1071 

87 

.158 

-578 

0.528 

0.730 

3-995 

.256 

3-433 

1872-75 

399,397 

n,584 

1066 

09 

.516 

.210 

o-437 

0.572 

2.736 

0.800 

2.342 

1876-80 

424,586 

17,876 

1147 

100 

.811 

.132 

o-3i7 

0.440 

2.709 

0.847 

2  .  306 

1881-85 

443,502 

16,688 

1025 

99 

.408 

.108 

o.  263 

0.532 

2.312 

0.848 

2  .007 

1886-90 

477>633 

26,654 

975 

117 

0.312 

I.OI5 

o.  196 

0-517 

2.042 

0.913 

I.  806 

1891-95 

57^463 

^0,804 

970 

123 

0.281 

0.806 

o  .  194 

0.934 

1.706 

0.822 

L531 

1896-03 

618,507 

50,404 

896 

125 

0.17 

0.764 

0.144 

o-457 

1.451 

O.SOI 

1.381 

Each  American  State  reveals  continuous  improvement  in 
the  underground  hazard;  but  for  the  entire  United  States  the 
death-rate  has  not  been  reduced.  The  reason  for  this  is  that 
new  coal-fields  and  new  mines  are  opened  from  time  to  time. 
The  new  difficulties  and  the  increase  of  unskilled  labor  naturally 
increase  the  rate  for  the  district.  This  is  apparent  in  the  new 
Western  coal  regions,  where  the  rate  is  greater  than  in  the  well- 
explored  Eastern  districts.  Indeed  it  is  inordinately  high. 

Discipline  in  Mines. — Unquestionably  a  more  rigid  discipline 
in  the  operations  of  mines,  a  general  education  of  the  men,  and  a 
wiser  discrimination  in  the  selection  of  workers  is  required,  and 
will  contribute  greatly  to  the  safety  of  mining  as  an  occupation. 
It  is  true  that  many  causes  of  accidents  are  unforeseen  and  un- 


4S8 


MANUAL  OF  MINING. 


avoidable,  but  it  is  also  true  that  the  largest  percentage  of  the 
causes  of  accidents  are  those  of  a  preventable  nature,  largely  in 
the  hands  of  the  miners  themselves.  If  a  code  of  rules  were 
officially  issued  dealing  with  the  severe  forms  of  mine  disaster, 
the  first  step  would  be  taken  toward  greater  security.  Again, 
the  present  laws  err  in  not  placing  direct  responsibility  for  defec- 
tive conditions  upon  some  one.  If  this  were  remedied,  better 
conditions  would  prevail.  Not  only  should  the  employers  be 

ACCIDENT  STATISTICS  OF  NORTH  AMERICA,  1893-1902. 


State. 

Deaths. 

Rate  per 

1000. 

Tons  Pro- 
duced per 
Death. 

Tonnage 
per 
Employee. 

353 
4i3 
8i3 
J45 
239 

212 
177 

118 

71 
135 
I°5 

540 
4344 
2218 
422 
237 
257 
911 
321 
119 

3.00 

5-59 

2.21 

2-57 
6.22 

2-34 
1.82 

i-35 
1.56 
1.76 
7.61 
1.76 
3.01 
2.30 
6.01 
24-75 
7-43 
3-73 
9-83 
1.96 

194,870 

110,000 

276,347 

365,518 

67,700 
221,849 
276,271 
360,000 

'61,366 

218,000 

83,362 

292,596 
126,795 
287,601 
71,118 
32,236 
69,232 
184,413 

68,138 

219,266 

500 

592 
607 

938 
421 
5i8 
378 
485 
957 
383 
634 
5H 
38i 
660 
437 
764 
5M 
687 
669 
433 

Illinois 

Kentucky  

Maryland  

Missouri  

New  Mexico  

Ohio                          

bituminous.  ....... 

Utah 

West  Virginia 

British  Columbia  

Nova  Scotia  

Total  North  America  

12,150 

2.89 

187,384 

54i 

fined  for  infringing  or  ignoring   the  laws,   but   the   employees 
should  be  penalized  for  hazarding  life  and  property  of  others. 

The  Causes  of  Accident. — The  following  prominent  causes 
of  underground  accidents  are  classified  according  to  frequency 
and  number  of  lives  involved.  Falls  of  roof  or  sides  of  rock  or 
coal;  accidents  in  the  haulage- ways;  injuries  in  shafts;  prema- 
ture explosions  of  powder  and  explosions  of  dust  and  gas.  The 
first  three  classes  of  accidents  rarely  cause  more  than  one  death 


ACCIDENTS  IN  MINES. 


489 


or  a  single  injury  at  a  time.     Explosions  of  gas  or  eruptions 
of  water,  fortunately  rare,  have  involved  hundreds. 

Falls  of  Ground. — This  form  of  accident  arises  from  the  fall 
of  roof  or  of  coal  in  working  places  and  in  roadways.  About 
70  per  cent  of  all  accidents  arise  in  the  rooms,  and  about  30  per 
cent  of  the  deaths  occur  in  roadways.  Each  accident  that  occurs 
involves  but  a  single  fatality  or  injury.  Yet  it  takes  the  lead  as 
the  most  prolific  cause  of  underground  fatalities.  The  grave 
danger  lies  in  the  fact  that  the  accident  is  local  and  of  continual 

FATAL  ACCIDENTS  IN  THE  PRINCIPAL  COAL-FIELDS  OF  NORTH  AMERICA. 


States:   1893  to  1902. 

Aggregate 
Number  of 
Persons. 

Total  Num- 
ber of  Lives 
Lost. 

Mortality 
per  1000. 

Tons  of 
Coal  per 
Death. 

Pennsylvania,  bituminous;    Ohio, 
Maryland  
Indiana,  Illinois,  West  Kentucky.  . 
Missouri,   Iowa,   Kansas,   Indian 
Territory  
Tennessee,   West   Virignia,    East 
Kentucky,  \labama. 

1,316,608 
456,341 

303.I47 

2,829 
1,017 

763 

2lII 
2.26 

2.52 

•i   70 

296,658 
293,874 

170,000 
16?  636 

Colorado,  New  Mexico,  Utah.  .  .  . 
Washington,  British  Columbia..  .  . 
Novia  Scotia,  1887-96  

97,296 
67,263 
60,716 

755 
578 
119 

7-77 
8-59 
1.96 

79,536 
68,245 
219,266 

Total  North  America,  bituminous  . 

2,758,756 

7,806 

2-73 

228,985 

Pennsylvania,  anthracite.  ....... 

1,443,110 

4,344 

3.01 

126,795 

Total  North  America  

4,201,866 

12,150 

2.89 

187,384 

occurrence  at  a  place  where  the  men  are  otherwise  engaged  at 
work,  and  ignore  or  fail  to  notice  the  warning  which  usually 
precedes  disaster.  The  weak  spots  in  the  roof,  the  various  fossil 
tree-trunks,  horses,  balls  of  ironstone,  rocks,  naturally  creviced, 
or  shaken  by  the  vibration  of  a  neighboring  blast,  are  liberated 
with  little  warning.  To  these,  with  the  coal  scaling  from  the 
sides  or  roof,  the  unnoticed  yielding  of  pillars  which  are  too  thin 
to  support  the  rock,  and  the  underholed  coal,  the  accidents  are 
due. 

Of  the  total  fatalities  in  the  anthracite  region  of  the  United 
States  during  the  past  twenty  years,  3521  were  caused  by  fall 


49° 


MANUAL   OF  MINING. 


of  rock  and  roof  in  the  working,  or  over  50  per  cent  of  the  total 
deaths  during  that  time.  In  bituminous  regions  the  proportion 
is  equally  large  during  the  same  years.  The  mine  inspector  for 
Illinois  reports  for  the  twenty  years  prior  to  1903,  a  death-list  of 

CLASSIFICATION  or  DEATHS  FOR  1902.     PRINCIPAL  CAUSES. 


Deaths  from 

Pennsylvania. 

Illinois. 

Ohio. 

Ken- 
tucky. 

Bituminous. 

Anthracite. 

Deaths. 

In- 
juries. 

Deaths. 

juries. 

Deaths. 

.  In- 
juries. 

Deaths. 

.In- 

juries. 

Deaths. 

Falls  
Cars     

223 
47 
126 
8 
14 

437 
239 

20 
41 
36 

Il6 
42 

20 

55 

256 

*33 

1? 

i°3 

59 
15 
4 
13 

203 
125 
15 

22 

54 

10 

3 
5 

162 

83 

3 
*3 

5 
5 

6 

Gas  
Powder  

'Surface  
Totals.  ... 

456 

861 

300 

641 

99 

406 

81 

298 

19 

THE  STATISTICS  OF  ACCIDENTS  IN  THE  PROMINENT  COAL-PRODUCING  COUNTRIES 
OF  EUROPE,  AND  Two  AMERICAN  STATES. 


Death-rate  of  Persons. 

fi 

: 

Countries. 

Periods. 

Persons 

Tons  of  Coal. 

II 

1^1 

( 

Q  2 

Pow- 

-j-a 

~"    In    ** 

Palls. 

Cars. 

o.  o 

der. 

Is 

H     ° 

Austria  

1894-03 

125,303 

367,088,336 

0.282 

.586 

.310 

.044 

1.718 

',937 

1896-03 

136  880 

i  786 

France  

1896-03 

i68,6oq 

243,671,927 

1335 

Great  Britain 

1896-04 

774,9" 

1,039,879,428 

o  713 

.I2S 

.114 

.042 

•3i 

8,100 

Prussia  

,  896^3 

443,995 

1,049,557,783 

0.98 

.68 

•*3 

.  II 

.16 

6,948 

Penn.,  anth  j 

1882-91 
1892-01 

1,059,526 
1,423,607 

351,058,927 
455-941,943 

'•37 
'•4.1 

.67 
•S9 

.29 
.27 

.27 

•30 

.22 
.  II 

4,427 

Penn.,  bit  

1893-02 

963,767 

638,213,809 

1.42 

34 

.24 

.06 

.26 

2,184 

Illinois  

1893-02 

368,607 

223,645,612 

1.28 

.21 

-38 

.08 

•25 

813 

1392,  of  which  nearly  50  per  cent  is  due  to  falls.  Doubtless  all 
over  the  United  States  the  number  of  lives  lost  from  this  cause 
bears  a  similar  proportion  to  the  total.  The  death-rate  from  falls 
of  roof  and  coal  alone  in  the  States  is  far  more  than  is  shown  by 
the  European  statistics  for  the  five  years  ending  with  the  year 
1900:  France,  0.58;  Great  Britain,  0.78;  Prussia,  1.22;  Illinois, 


ACCIDENTS  IN   MINES.  491 

1.34;  Pennsylvania  (bituminous),  1.84;  and  Pennsylvania  (an- 
thracite), 2.1 1.  In  Great  Britain  22,190  lives,  or  46.5  per  cent/ 
were  lost  by  falls  since  1853. 

The  death-rate  is  large  and  is  not  decreasing  proportionately 
with  other  causes,  for  the  conditions  in  the  working  places 
depend  entirely  upon  the  miner  himself,  and  here  improvement 
is  not  manifested.  It  is  true  that  the  pressure  of  the  roof,  and 
the  dislodgement  of  the  fragments  therefrom,  cannot  be  indefi- 
nitely resisted  or  avoided,  but  its  movement  can  be  detected  and 
checked  by  systematic  timbering.  The  best  means  of  detection 
is  the  use,  near  the  face,  of  wooden  props,  not 'iron,  whose  bend- 
ing will  give  warning  of  disaster  by  buckling.  A  better  illumina- 
tion will  reveal  the  condition  of  the  roof,  continual  testing  would 
disclose  its  loose  fragments,  and  a  systematic  timbering  would 
materially  diminish  the  risk  of  its  fall.  The  use  of  coal-cutting 
machines  would  remove,  the  dangers  of  underholing  coal;  and 
so  would  the  enforcement  of  a  specific  order  for  the  dismissal  of 
any  employee  failing  to  utilize  the  props.  An  accident  is  the 
result  of  deliberate  neglect,  and  the  delinquent  is  more  frequently 
the  old  hand  than  the  newcomer. 

Rigid  supervision  and  the  appointment  of  a  suitable  timber- 
boss  to  visit  all  the  working  places  and  regulate  the  timbering, 
with  power  to  punish  carelessness,  would  be  a  most  effective 
remedial  measure.  In  certain  districts  in  France  the  rigid  en- 
forcement of  regulations  requiring  systematic  timbering  imme- 
diately upon  advancing  beyond  the  distance  fixed  by  the  regu- 
lation has  reduced  the  loss  of  life  to  about  |  man  per  1000 
from  the  previous  rate  of  about  i  in  1000.  In  1870-1879  it  was 
0.90  per  1000;  to  1886,  0.24  per  1000;  to  1890,  0.15  per  1000; 
to  1892,  0.13  per  1000.  In  Great  Britain  the  rate  of  decrease 
during  the  same  period  from  1.20  to  0.764,  though  a  decided 
improvement,  was  not  so  marked. 

Accidents  Due  to  Cars,  Shafts,  etc. — Second  on  the  list  of 
accidents  in  all  classes  of  mines  are  those  occurring  in  the  trav- 
elling-ways. The  men  fall  from,  or  are  run  over,  by  cars.  Such 
accidents  are  more  numerous  in  coal-mines  than  in  metal-mines, 


-492  MANUAL  OF   MIXING. 

because  the  speed  of  the  cars,  or  trains,  and  the  number  of  men 
employed  are  greater.  Many  casualties  arise  from  jumping  on 
or  off  the  trains  while  in  motion.  These  are  manifestly  invited 
by  the  victims  themselves;  other  than  these,  the  accidents  along 
the  travelling-ways  are  those  arising  from  runaway  cars,  or 
the  lack  of  clearance  in  haulage-ways.  Numerous  safety-niches 
.along  the  line  will  afford  the  men  some  protection.  Those 
occurring  in  the  shaft  are  largely  averted  by  a  security  of  the 
hoisting  appliances,  an  efficient  system  of  signals,  and  the  use 
of  overwinding  and  safety  attachments.  An  average  of  5  deaths 
from  overwinding  occurred  annually  in  Great  Britain.  Cases 
in  which  men  are  caught  between  the  cage  and  the  shaft-timbers 
-or  are  struck  by  material  falling  from  above  are  purely  accidental 
and  may  be  classed  as  inherent  to  the  occupation.  Those  occur- 
ring from  the  fainting  of  overheated  men  while  being  hoisted 
from  the  shaft  are  not  numerous.  These  can  be  prevented  by 
gates  at  the  sides  of  the  cages  and  hoods  overhead.  The  Euro- 
pean shafts  are  more  fruitful  of  accident  than  the  shallow  Amer- 
ican shafts. 

Accidents  from  Use  of  Explosives. — In  metal-mines  the  per- 
centage of  injuries  due  to  premature  blasts  is  larger  than  in  coal- 
mines. They  are  more  .frequently  the  result  of  carelessness 
.arising  from  the  use  of  iron  bars  while  loading  a  charge,  from 
tampering  with  the  metallic  caps,  handling  powder  near  an 
exposed  flame,  thawing  out  frozen  dynamite,  drilling  into  unex- 
ploded  cartridges,  attempting  to  ignite  too  many  blasts  at  one 
time,  and  returning  to  the  work  too  soon.  These  accidents  are 
manifestly  local  in  their  nature  and  occur  either  at  the  working- 
face,  or  at  the  powder-house  where  the  explosive  is  stored.  The 
victims  themselves  are  the  active  cause.  The  comparatively 
high  death-rate  in  the  anthracite  coal-mines  is  attributed  by 
some  inspectors  to  the  great  consumption  of  high  explosives. 

The  use  of  electric  cartridges  and  the  prohibition  of  loose 
powder  for  firing  in  dusty  collieries  would  materially  diminish 
the  number  of  casualties.  Greater  care  in  the  selection  of  the 
•explosive  and  its  use  can  readily  be  exercised  by  both  the  em- 


ACCIDENTS  IN  MINES.  493 

ployers  and  the  employees.  In  some  mines  blasting  is  assigned 
to  a  special  class  of  men. 

Gas  Explosions. — No  coal-mine  is  free  from  gas,  and  no  col- 
liery can  be  operated  without  illuminating  and  blasting  ajents. 
Every  mine,  therefore,  is  subject  to  ignition  of  a  greater  or  less 
volume  of  gas.  The  result  may  be  a  fire  or  an  explosion,  accord- 
ing as  the  amount  of  air  present  be  excessive  or  at  the  critical 
percentage,  and  its  effect  will  be  proportional  to  the  volume  of  gas 
or  the  quantity  of  coal-dust  present.  All  dry  mines  should  be 
ranked  as  fiery  whether  gaseous  to  a  dangerous  degree  or  not. 

A  steady  flow  of  escaping  gas  into  the  regular  working  places 
rarely  results  in  dangerous  conditions,  for  the  volume  of  the  air- 
current  there  dilutes  it  to  an  innocuous  degree.  But  when  some 
internal  reservoir  of  gas  bursts  into  the  entries  or  abandoned 
rooms,  where  the  circulation  is  sluggish,  or  the  air  comparatively 
small  in  volume,  the  mixture  of  gas  and  air  may  easily  become 
critical.  This  danger  is  imminent  in  the  driving  of  galleries 
far  in  advance  of  faces  in  the  rooms  or  through  a  seam  beyond  a 
fault  where  heavy  volumes  of  gas  may  be  liberated,  particularly 
if  the  ground  slip  or  be  soft.  A  blast  or  a  flame  in  contact 
with  the  mixture  may  ignite  it  and  cause  disaster.  One  gas 
explosion  involving  62  lives  occurred  within  twenty-four  hours 
after  the  visit  to  a  mine  regarded  by  the  chief  inspector  as  "the 
safest  and  best-conducted  mine  in  the  State,"  but  caused  by 
striking  a  clay-seam.  The  ignition  came  from  either  a  defect- 
ive lamp  or  an  open  lamp.  This  is  usually  the  direct  cause. 
Men  will  disregard  the  prohibitory  notice  of  the  fire-boss  barring 
them  from  a  given  working  place,  and  enter.  For  one  cause  or 
another  they  will  uncover  the  flame  of  their  lamp.  The  reports 
for  Great  Britain  show  that  out  of  the  total  of  gas  or  dust  explo- 
sions about  70  per  cent  owe  their  injuries  to  the  use  of  naked  lights 
or  imperfect  lamps;  18  per  cent  to  shot-firing,  and  14  per  cent 
to  the  accidental,  or  spontaneous,  ignition  of  the  mineral.  This 
form  of  accidents  during  the  last  twenty  years  has  caused  1451 
deaths,  or  0.35  out  of  every  1000  coal-mines  in  Pennsylvania. 
It  was  9  per  cent  of  the  total  fatalities.  Explosions  killed  over 


494  MANUAL  OF  MINING. 

10,000  men  during  the  past  fifty  years,  but  the  rate  is  decreasing 
very  much,  as  shown  in  Great  Britain  for  the  successive  semi- 
decades  between  1852  and  1003. 

R2medial  Measures. — The  measures  which  will  safeguard  the 
men  from  the  consequence  of  sudden  outbursts  of  gas  include, 
besides  the  dilution  of  the  gas  in  wide  galleries,  the  use  of 
numerous  bore-holes  to  indicate  an  approach  to  a  gaseous 
mass.  This  practice  is  sharply  criticised  by  many  because  of 
the  difficulty  of  boring  in  shattered,  and  consequently  infested, 
portions  of  the  seam.  Some  mines  dispose  of  the  gas  through 
bore-holes,  by  which  it  is  discharged  to  the  surface  and  burned. 
An  increased  volume  of  air,  its  better  distribution  to  working- 
faces,  and  the  discontinuance  of  heavy  blasts  of  explosives  w?hile 
the  men  remain  in  the  mine  are  diminishing  the  number  of  in- 
juries and  fatalities  from  explosions.  There  remains  much  to 
be  done  by  the  universal  employment  of  some  bonneted  lamp 
which  is  safe  and  at  the  same  time  furnishes  good  light. 

A  number  of  deaths  charged  to  explosions  occur  in  the  exits 
of  the  mine  where  the  fleeing  victims  have  been  overwhelmed  by 
after-damp.  Rescuing  parties  find  them  generally  in  the  intake. 
Here  it  is  that  the  greatest  force  of  the  concussion  is  exercised, 
and  along  it  too,  the  gases  make  their  exit.  The  return  airway 
is  unquestionably  far  safer  than  is  the  intake  airway  after  an 
explosion,  particularly  if  solid  air-crossings  exist  in  the  mine. 
Many  cou  d  have  made  their  escape  through  that  wray  had  they 
been  accustomed  to  this  outlet. 

The  Theory  of  Coal-dust  Explosions. — The  circumstances  of 
many  explosions,  particularly  of  those  on  a  large  scale,  cannot 
be  explained  fully  by  reference  to  gas  alone.  Those  mines  in 
which  explosions  occur,  and  are  known  as  fiery,  are  invariably 
dry.  These  dry  mines  are  dusty,  containing  a  large  quantity  of 
the  mineral  charcoal,  "mother  of  coal,"  which  is  continually 
afloat  in  the  atmosphere.  The  explosions  being  more  frequent 
in  dry  mines  and  deep  mines  indicate  the  influence  of  the  coal- 
dust  in  extending  and  aggravating  the  danger. 

Different  fine  dusts  are  inflammable  and  consequently  dan- 


ACCIDENTS  IN   MINES.  495 

gerous,  according  to  their  degree  of  fineness  and  chemical 
composition.  Lycopodium-powder,  which  is  a  modern  vege- 
table product  resembling  mineral  charcoal,  is  highly  explosive. 
Hence  it  may  be  a  cause,  and  even  the  sole  cause,  of  the 
explosion.  It  cannot  be  ignited  except  by  direct  contact  with 
an  intensely  hot  flame.  There  is  no  probability  that  the  ordinary 
flame  of  a  lamp  has  produced  explosion  from  coal-dust  alone, 
nor  has  it  been  shown  that  the  ord  nary  blown-out  shot  has 
ignited  coal-dust  without  the  presence  of  some  gas.  The  danger 
of  explosion  does  exist  when  both  gas  and  coal-dust  are  present. 
It  is  able  to  extend  indefinitely  the  transmission  of  an  explosive 
flame  and  thus  intensify  the  shock.  Experiments  have  proven 
that,  without  any  floating  dust,  the  flame  from  a  blown-out  shot 
would  not  travel  more  than  25  feet,  but  that  soot  would  convey 
the  flame  to  200  feet  The  conclus  on  may,  therefore,  be  drawn 
that,  though  coal-dust  alone  may  not  be  dangerous,  in  the  pres- 
ence of  gas,  even  in  small  quantities,  it  becomes  highly  so.  • 

A  preventative  remedy  consists  in  laying  the  dust  by  sprink- 
ling. A  spray  can  be  delivered  into  the  air- current  from  pipes 
i  inch  or  2  inches  in  diameter  under  pressure  of  50  Ibs.  per 
square  inch  to  moisten  the  air  and  materially  assist  in  the  decom- 
position of  the  dust 

Ankylostomiasis. — This  is  a  disease  attacking  miners  in  the 
wet  collieries,  and  though  but  recently  discovered  in  Germany, 
is  receiving  considerable  attention  from  other  governments 
because  of  its  infectious  character,  and  energetic  measures  are 
being  taken  to  eradicate  it  Imported  into  Westphalia  from 
Hungary,  in  one  district  alone  it  has  increased  from  107  cases 
in  1896  to  1400  cases  in  1902,  and  during  the  year  1902,  in  another 
district  of  Germany,  5.29  miners  out  of  every  1000  were  afflicted. 

The  disease  resembles  cholera  or  typhoid,  the  source  of  con- 
tagion being  faecal  matter.  It  is  treated  by  the  medical  men 
as  dangerous  and  the  patients  are  quarantined.  It  is  promoted 
by  the  sprinkling  of  the  mine  with  pit-water,  and  by  the  high 
underground  temperature.  Inasmuch  as  dry  mines  of  the 
infected  districts  do  not  show  so  large  a  number  of  cases,  un- 


496  MANUAL  OF  MINING. 

doubtedly  the  discontinuance  of  watering  for  a  time  would  pre- 
vent the  development  of  the  ova.  The  obstacle  to  this  method 
is  the  increased  danger  from  explosions  from  gas  or  dust. 
The  nature  of  the  disease  is  not  perfectly  understood  as  yet, 
and  many  miners  have  been  treated  two  or  three  times  without 
the  removal  of  the  parasite.  Its  destruction  in  human  excre- 
ment is  best  effected  by  the  use  of  fern  extract  (Extractum  Filicis), 
which  proves  a  better  remedy  than  the  thymol  used  in  England, 
Dry-earth  pails,  one  to  every  four  workers,  are  distributed  to 
prevent  dissemination.  The  mine  owners  are  aware  of  the 
gravity  of  the  situation,  and  in  many  districts  have  discontinued 
the  watering  to  reduce  the  plague.  That  it  will  in  time  develop 
in  our  American  mines,  among  the  foreign  element  coming  from 
the  infected  regions,  there  is  little  doubt,  and  careful  watch  should 
be  kept  for  its  appearance. 

Underground  Fires. — The  most  extensive  conflagrations  in 
mines  are  those  initiated  by  gas  or  powder  explosions,  com- 
municated subsequently  to  timbers,  canvas,  or  floating  coal-dust. 
Fires  differ  from  explosions  only  in  the  degree  of  combustion,  and 
either  may  result  in  the  other.  The  flame  from  a  miner's  lamp 
or  a  spark  from  powder,  in  contact  with  any  combustible, 
are  the  elements  producing  explosions  or  fire,  and  the  remedy 
is  to  isolate  one  or  all  of  them  The  exercise  of  care  in 
stables,  pump-rooms,  and  storage-bins  for  oil,  powder,  and  waste 
would  eliminate  a  grave  danger.  In  these  quarters  electric 
lights  can  well  be  provided  and  the  floors  sprinkled  with  clean 
gravel  or  sand.  In  rooms  or  travelling- ways  fires  come  from 
defective  lamps  or  exposed  lights. 

Fires  from  Blasting  Agents. — Powder  is  not  liable  to  ignite  fire- 
damp because  the  temperature  of  its  detonation  is  not  over  1000° 
F.  An  explosive  not  properly  confined  to  the  hole  in  which  the 
execution  is  to  be  done  will  be  blown  out  in  the  room  and  ignite 
any  combustible  gas,  if  sufficient  air  be  present.  Blasting 
agents  which  are  not  chemically  perfect  may  also  produce  fire 
and  explosion.  The  unconsumed  gases  of  the  detonator  are 
projected  into  the  workings,  where  they  combine  with  the  requisite 


ACCIDENTS  IN   MINES.  49  T 

amount  of  oxygen  to   cause   a   second  blast,   under  conditions- 
dangerous  to  life  and  property. 

The .  employment  of  a  better  grade  of  explosive,  the  use  of- 
a  powder  free  from  flame,  the  ignition  of  the  fuse  without  spark,, 
the  use  of  water  cartridges  for  extinguishing  the  flame,  the  em- 
ployment of  specialists  to  whom  the  blast  is  entrusted,  and  the- 
substitution  of  electric  firing  when  the  employees  have  left  the 
works,  will  lessen  the  risk. 

Spontaneous  Combustion. — This  is  the  slow  process  of  burn- 
ing where  heat,  combustibles,  and  the  minimum  amount  of  air' 
are  the  elements  involved,  .and  it  frequently  occurs  in  a  goaf 
and  in  abandoned  rooms  containing  broken  coal,  dust,  and  rotting' 
timber.  Here  the  roof  pressure  compacts  the  material,  develop- 
ing an  amount  of  heat  sufficient  to  partially  decompose  the  ever- 
present  pyrites.  This  incipient  conflagration  is  furthered  by 
ingress  of  air.  The  amount  of  combustible  consumed  depends 
upon  the  amount  of  air  gaining  access  to  it.  Once  started,  a. 
pressure  ensues  within  the  room  or  goaf,  and  "breathing"  follows, 
forcing  gas  out  through  the  trenches  in  the  bottom,  at  the  roof, 
or  from  cracks  in  the  side  walls  and  admitting  air.  At  this- 
point  the  sulphurous  fumes  give  warning,  and  active  measures, 
must  be  taken  to  prevent  fire  or  explosion,  or  both.  But  a 
complete  isolation  from  the  mine  is  difficult  and  a  perfectly 
impermeable  enclosure  impossible,  hence  the  failures  attending; 
the  employment  of  these  measures.  The  safer  method  of  pro- 
cedure would  be  the  construction  of  a  gas-tight  bulkhead,  or 
dam,  against  the  goaves,  if  they  can  be  perfectly  walled.  If  not, 
their  contents  should  be  removed,  though  this  can  be  done  only 
at  great  risk. 

Spontaneous  combustion  can  be  avoided  by  keeping  cool  the- 
combustible  accumulations  in  the  rooms  or  abandoned  work- 
ings by  an  active  air- current. 

Extinguishing  Fires. — Besides  the  remedies  already  sug- 
gested for  preventing  their  inception,  the  usual  method  of  pre- 
venting the  spread  of  fire  is  to  beat  out  the  flames  or  drown 
the  fire  with  a  flow  of  water,  steam,  or  carbonic  acid.  Resort: 


49 8  MANUAL  OF  MINING. 

must  be  had  to  the  last  method  when  the  fire  cannot  be  extin- 
guished by  cruder  methods,  or  when  it  has  attained  such  head- 
way as  not  to  be  overcome  by  simple  means.  To  be  effective, 
however,  the  heading  must  be  absolutely  gas-tight,  not  only  in 
the  bulkheads  but  also  in  the  upper  strata.  If  they  are  porous  or 
the  mine  is  shallow,  air  will  enter  from  above  to  feed  the  flame. 
The  mine  in  that  event  must  be  abandoned.  A  fire  at  the  Calu- 
met and  Hecla  copper-mines,  said  to  have  been  communicated 
to  the  shaft-timbers  by  the  friction  of  the  hoist-rope  rollers,  was 
extinguished  by  the  liberal  use  of  carbonic  acid.  After  the  surface 
had  been  frozen  to  stop  leaks  in  the  shaft,  carbonic  acid  was 
injected  and  the  fire  brought  under  control.  For  the  manufacture 
of  3000  cubic  feet  of  carbonic  acid  there  were  used  1200  gallons 
of  sulphuric  acid  and  4500  Ibs.  of  limestone. 

Water  is  the  simplest  quencher  of  flames,  provided  it  can  be 
made  to  reach  all  portions  which  are  on  fire.  But  it  has  happened 
that  into  certain  rooms  the  water  above  the  foot  of  the  shaft 
has  not  reached  the  face  because  of  the  compression  of  the  air 
locked  in  them,  which  cannot  escape.  The  fire,  therefore,  raged 
above  the  water  level  until  this  air  was  consumed,  and  in  some 
cases  broke  out  again  as  soon  as  the  water  was  removed. 

Steam  has  been  injected  into  some  burning  rooms  for  quench- 
ing fires,  but  it  has  not  proven  very  acceptable. 

Protective  Measures. — As  a  prevention  against  the  excite- 
ment and  confusion  arising  at  the  time  of  an  explosion  or  a  fire, 
the  following  disciplinary  precautions  may  be  exercised  before 
the  accident: 

i.  Consider  the  quickest  and  safest  mode  of  descending  into 
the  mine  when  the  usual  winding  arrangements  are  useless. 
2.  Plan  for  the  installation  of  a  special  winding-engine.  3.  Con- 
nect water-pipes  from  the  surface  to  the  mine  with  branches 
for  use  in  case  of  fire.  4.  Arrange  for  the  fitting  of  an  extra 
engine  and  fan  for  the  emergency.  5.  Keep  the  tracings  of 
all  working  plans  to  within  three  months  of  work,  showing  all 
roads  then  open,  the  position  of  overcast,  doors,  and  brattices. 
6.  Accustom  the  men  periodically  to  travel  certain  roads  which 


ACCIDENTS  IN  MINES,  499 

they  are  not  in  the  habit  of  taking,  having  finger-boards  showing 
the  direction  of  the  upcast  pit.  7.  Keep  on  hand  all  safety 
apparatus,  including  the  Fleuss  machine,  a  quantity  of  light  air- 
pipes,  and  "first-aid"  appliances.  8.  Appoint,  during  ordinary 
working  of  the  collieries,  some  of  the  leading  officials  to  act  as 
emergency  officers  at  the  time  of  accident,  drilling  them  in  their 
duties. 

Rules  for  Guidance  after  Explosions. — The  following  mode 
of  procedure  would  facilitate  the  work  of  the  rescuing  parties 
as  presented  by  Mr.  Garforth.  Instructions  similar  to  these 
should  be  posted  about  the  mine,  in  various  languages,  for  the 
guidance  of  the  men  and  the  emergency  officers.  Each  colliery 
and  mine  should  have  a  complete  map  of  its  workings  and  a 
perfect  outline  indicating  correctly  the  path  or  paths  pursued 
by  the  ventilating  current  in  its  circuit  through  the  mine.  It 
is  necessary  when  accident  occurs  as  a  guide  for  the  rescuing 
parties,  and,  although  the  mining  laws  do  not  require  this  pro- 
vision, it  would  be  well  if  maps  were  quickly  available. 

i.  Send  for  the  emergency  officers,  assign  to  each  his  duty, 
and  appoint  some  one  as  deputy  in  the  event  of  serious  accident 
to  the  manager.  2.  Examine  all  old  connections  with  the  shaft 
and  arrange  to  repair  broken  stoppings.  Prepare  stretchers  and 
stimulants  and  arrange  for  a  hospital.  3.  Provide  exploration 
parties  of  five  with  leaders,  supplied  with  safety-lamps,  mine- 
plans,  restoratives,  cylinders  of  oxygen,  and  a  stout  cord.  On 
no  account  must  any  one  enter  alone,  even  on  the  shortest  jour- 
ney. 4.  In  advancing,  the  party  should  move  in  single  file,  the 
leader  of  each  search-party  alone  testing  for  gas.  Do  not  let  a 
safety-lamp  be  the  final  guide  as  to  the  absence  of  after-damp. 

5.  Loss  of  life  to  explorers  may,  perhaps,  be  avoided  by  remem- 
bering the  dangers:   (a)  after-damp;    (6)  falls  of  roofs  and  sides; 
(c)  underground  fires  and  consequent  risk  of  a  second  explosion. 

6.  If  the  force  of  the  explosion,  has  blown  out  the  separation- 
doors  and  overcast,  they  should  not  be  restored  because  of  the 
possibility  of  undiscovered  fire.     7.  Main  intake  airways,  blocked 
by  falls,  must  not  be  traversed  unless  the  men  carry  with  them 


5oo 


MANUAL  OF  :.II:-:I::G. 


an  unrolled  brattice-cloth,  which  will  admit  of  a  double  current  of 
air'  through  the  airway.  The  brattice  should  be  non-inflam- 
mable. 8.  To  discover  the  existence  of  fire,  restore  the  ventila- 
tion and  examine  the  return  airways  every  hour  for  (a)  fire- 


FIG.  171. — A  Diving  Knapsack. 

stink  and  (ft)  a  rise  in  temperature.  If  the  former  be  noticed, 
that  section  from  which  it  comes  should  be  explored  first,  and 
its  fire  extinguished  if  possible,  t>r  it  should  be  closed  off  by  stop- 
pings, or,  in  extreme  cases,  the  pit  entirely  closed.  9.  Parties 
should  be  careful  not  to  go  too  far  at  once,  even  when  taking  air 
with  them,  as  the  force  of  the  explosion  will  have  forced  the 


ACCIDENTS  IN  MINES.  501 

after-damp  into  the  interstices  of  the  goaves,  whence  it  will 
gradually  exude. 

Aerophores. — For  penetrating  a  very  impure  atmosphere 
aerophores  of  different  makes  are  to  be  had.  They  consist  of  a 
portable  bag  or  cylinder  carrying  enough  compressed  air  or  oxy- 
gen for  the  respiration  of  a  miner,  and  his  lamp,  while  making 
repairs  or  exploring.  The  oxygen  is  inhaled  by  one  tube,  while 
through  an  exhaler  is  ejected  the  CO2,  which  is  absorbed  by 
caustic  soda,  leaving  the  N  only  to  return  to  the  bag.  Fleuss' 
apparatus  looks  like  a  knapsack,  weighs  28  Ibs.,  contains  a  four- 
hours'  supply  of  oxygen,  and  has  besides  a  self-contained  illumi- 
nator— an  acetylene  lamp  does  not  depend  upon  the  oxygen  of 
the  air  for  its  burning.  A  lamp  burning  methylated  spirit  heats 
a  plug  of  lime  and  renders  it  incandescent. 

The  Fleuss  diving  knapsack  (Fig.  171)  consists  of  a  cylinder 
and  a  cell  in  four  compartments  with  a  perforated  false  bottom. 
The  cylinder  contains  oxygen  at  240  Ibs.  pressure,  and  delivers 
the  gas  to  the  nostrils  by  a  tube.  The  carbonic  acid  gas  is  ex- 
haled by  the  diver  into  the  cell,  where  it  is  absorbed  by  caustic 
soda.  The  entire  combination  carries  a  four-hour  supply,  and 
has  done  excellent  service  to  rescuing-parties  after  accidents 
arising  from  fire  and  inbursts  of  water  or  floods  of  gas. 


REFERENCES. 

Safeguards  against  Danger  from  Mine  Gases,  Am.  Mfr.,  Jan.  15,  1897,  83; 
Miners'  Phthisis,  William  Cullen,  Jour.  Chem.  Met.  Soc.  of  S.  Africa,  March 
1903  and  Feb.  1903;  Gases  in  Metalliferous  Mines,  M.  &  M.,  Vol.  XXIV, 
478. 

First  Aid  the  Injured,  M.  &  M.,  Vol.  XXIII,  151. 

Accident  and  Skeleton  Maps,  M.  &  M.,  Vol.  XX,  115. 

Damping  the  Air  of  Coal  Mines  as  a  Safeguard  against  Explosions,  James 
Ashworth,  Ir.  &  Coal  Trds.  Rev.,  Sept.  26,  1902;  Spraying  Dust  Mines, 
M.  &  M.,  Vol.  XXIV,  306. 

Underground  Fires,  Lecture,  Coll.  Mgr.,  1894,  35;  Gob  Fires,  111.  Min. 
Inst.,  II,  1896;  Fire,  Min.  Insp.,  Pa.,  1886,  Packet;  Shaft  Fire  and  its  Lesson: 
Coll.  Guard.,  Oct.  30,  1896,  838;  Fires  in  Coal  Mines,  J.  Abadie,  Genie 
Civil,  Oct.  4,  1902;  Mine  Fires,  M.  &  M.,  Vol.  XXIII,  305;  The  Handling 


502  MANUAL  OF  MINING. 

of  Underground  Fires  with  Milk  of  Lime,  Wolfgang  Kummer,  Oesterr. 
Zeitschr.  f.  Berg.  u.  Huttenwesen,  April  9,  1904. 

Explosions,  the  Physics  of,  Coll.  Mgr.,  Jan.  18,  1895,  14;  On  Explosion 
to  Managers,  Precautions  Before,  Coll.  Guard.,  June  1897,  1084;  Dust  Ex- 
plosions, 111.  Min.  Inst.,  II,  40. 

Abandoned  Mines,  Protection  of,  Coll.  Mgr.,  Dec.  18,  1896,  662. 

Falls,  Accidents:  Scope  of  Committee  Inquiry,  Coll.  Guard.,  June  18,  1897, 
1186;  Recklessness  of  Men  with  Lamps,  C.  Le  Neve  Foster,  Coll.  Guard., 
Dec.  n,  1896,  1186;  Falls  of  Roof  and  Sides:  Accidents  and  Their  Pre- 
ventions, H.  M.  Mine  Insp.,  Coll.  Mgr.,  1896,  22;  An  Analysis  of  the 
Casualties  in  the  Anthracite-coal  Mines  from  1871  to  1880,  H.  M.  Chance, 
Amer.  Inst.  M.  E.,  X,  67;  The  Prevention  cf  Accidents  in  Mines,  Austin 
Kirkup,  Trans,  of  the  N.  of  Eng.  Inst.  M.  &  M.  Eng.,XLV,  Part  I,  2;  Com- 
pilation of  the  Fatal  Disabling  Accidents  in  the  Breslau  Mining  District,  Ge.- 
many,  Gluckauf,  April  7,  1900. 

Wages  and  Profit  in  Coal  Mining,  Ir.  &  Coal  Trds.  Rev.,  June  19,  1903. 

Prevention  of  Shaft  Accidents,  Coll.  Gua  d.,  LXXIX,  no;  Fire  Damp  and 
Relation  to  Internal  Meteorology,  Coll.  Guard.,  Vol.  LXXIX,  612;  Rescue 
Work,  Coll.  Guard.,  Vol.  LXXXII,  1159;  Diagram  of  Ventilation,  Coll. 
Guard.,  Vol.  LXXXII,  645;  The  Nature  of  Explosions,  Coll.  Guard.,  Vol. 
LXXXII,  1267;  Watering  Rooms,  Coll.  Guard.,  Vol.  LXXXIII,  1045;  Elec- 
tric Accidents  in  American  Mines,  Coll.  Guard.,  Vol.  LXXXIV,  150;  Anky- 
lostomiasis,  Coll.  Guard.,  Vol.  LXXXVI,  774;  Ambulances,  Coll.  Guard., 
Vol.  LXXXVI,  1346,  and  LXXXV,  464;  Watering  Roads,  Coll.  Guard.,  Vol. 
LXXXVI,  893. 


II. 

PRACTICAL  MINING. 

CHAPTER  I. 

SHAFTS. 

Shafts  may  be  sunk  for  permanent  or  temporary  purposes, 
and  they  may  be  intended  for  one  especial  purpose  only — of 
hoisting,  travelling,  or  ventilation;  or  their  size  may  be  suffi- 
ciently large  to  warrant  division  into  a  number  of  compartments, 
one  for  the  pumping  and  ladderway  and  the  remainder  for 
hoists,  according  to  the  output.  Collieries  require  additional 
communication  with  the  surface  for  ventilation.  The  large  area 
required  for,  and  the  foulness  of,  the  return  air  demand  a  sepa- 
rate outlet  for  the  upcast,  as  also  for  the  intake,  which  should 
never  be  interfered  with  by  hoisting. 

The  numerous  drawbacks  to  the  single-entry  compartment 
shaft  or  slope  are  so  fully  recited,  page  20,  that  only  in  vein-mines 
should  the  development  be  thus  risked.  Certainly  the  ventilat- 
ing-ways  should  not  be  in  adjoining  compartments,  because  the 
bratticing  could  never  be  kept  tight  enough  to  prevent  a  leakage 
of  fresh  air  into  the  upcast.  Only  unusually  hard  rock,  or  ex- 
ceptional difficulties  in  soft  or  watery  ground,  warrant  a  single 
entry.  Where  a  prospecting  drill-hole  has  been  used  to  test  the 
ground  the  shaft  should  not  be  carried  down  along  on  it.  It 
could  eventually  be  of  greater  service  as  a  ventilator  and  a 
ropeway  than  it  could  be  capable  of  during  sinking. 

503 


504  MANUAL  OF  MINING. 

When  it  is  desired  to  remove  the  mineral  quickly,  several 
shafts  are  sunk,  their  positions  being  a  matter  of  indifference. 
Ordinarily,  however,  the  location  of  a  shaft  and  its  equipment 
is  a  matter  of  vital  import.  The  configuration  or  nature  of  the 
surface  affecting  transportation  may  govern  the  selection  of  a 
site;  but,  cateris  paribus,  the  principal  shaft  should  be  so  located 
as  to  reach  the  lowest  point  of  the  workings.  This  is  not  at  the 
outset  always  possible  to  do  so  we  are  accustomed  to  see  one 
shaft  after  another  abandoned  or  relegated  to  secondary  uses. 
Instance  the  numerous  illustrations  from  the  Lake  Superior 
region.  The  Calumet  and  Hecla  has  eight  shafts,  each  over 
i coo  feet  deep,  with  a  complete  plant  over  each  one.  Nor  is 
this  an  exceptional  case. 

How  Deep  can  We  Mine? — This  question  is  not  now  the 
serious  one  that  it  was  twenty  years  ago.  The  natural  limits 
as  determined  by  physical  conditions  vary  Undoubtedly  the 
maximum  depth  in  the  copper-mines  of  Lake  Superior  will 
always  be  greater  than  in  porphyry  districts.  The  question  of 
ultimate  depth  would  be  a  limit  approachable  by  the  possibility 
of  installing  at  the  surface  sufficiently  large  hoisting-engines  and 
stout  enough  wire  ropes  to  lift  the  live  load  and  the  weight  of 
the  rope  itself.  These  mechanical  difficulties  have  a  positive 
limit  beyond  which  no  mining  can  be  carried. 

Again,  the  possible  reward  of  mining  compared  with  the 
expense  of  operations  would  place  a  limit  of  depth  which  is  inde- 
pendent of  the  mechanical  question.  Judging  from  the  cost  of 
mining  at  the  increasing  average  depth  in  copper-mines,  it  must 
be  granted  that  there  appears  to  be  no  fixed  relation  between 
depth  and  running  expense.  The  main  expenses  of  mining, 
such  as  stoping,  tramming,  and  superintendence,  are  to  some 
degree  fixed,  but  when  the  heat  begins  to  be  troublesome  with 
increased  depth,  the  cost  of  the  working  is  increased.  Pumping 
need  not  be  a  serious  element  in  determining  the  question  of 
cost,  for  experience  shows  in  the  deep  mines  a  proportionately 
smaller  quantity  of  water,  and  a  lower  price  per  ton  of  output, 
than  existed  during  the  earlier  history  of  the  mines  Increased 


SHAFTS.  505 

depth  means  increased  rock  pressure,  and  necessarily  stronger 
braces  in  shafts  and  all  openings. 

On  the  other  hand,  the  increasing  output  and  the  general 
improvement  in  mechanical  and  engineering  appliances  have 
resulted  in  a  saving  much  greater  than  the  increased  cost  due 
to  increased  depth.  Undoubtedly,  at  the  present  time,  with 
an  estimated  depth  of  mine  twice  that  of  fifteen  years  ago,  the 
total  cost  of  construction  is  less  per  ton  of  output.  Again,  the 
increased  depth  results  in  increased  capacity  of  the  mine,  which 
should  be  followed  by  a  corresponding  increase  in  the  capacity  of 
the  shaft  and  stoping,  and  by  a  change  in  the  methods  of  hoisting. 
Instead  of  a  single  compartment  with  the  intermittent  hoist  of  a 
single  cage,  there  must  be  a  number  of  skips  or  cages  in  each  of 
several  compartments.  This  condition  must  limit  the  depth  to 
6000  or  7000  feet. 

The  results  of  the  general  discussion  indicate  that  the  limit 
might  be  placed  in  the  Transvaal  or  our  Michigan  copper- 
mines  at  about  8000  feet,  unless  a  discovery  of  exceedingly  rich 
ore-deposits  below  this  depth  should  warrant  exceptional  appli- 
ances for  their  recovery. 

Some  of  the  deep  mines  and  shafts  of  the  world  are  men- 
tioned below: 

In  Austria-Hungary  7  are  deeper  than  1500  feet;  in  Bel- 
gium 12  exceed  2000  feet  in  depth;  in  France  5  are  over  2000 
feet;  Germany  has  10  over  2000  feet;  in  Great  Britain  are 
probably  100  exceeding  1500  feet  and  20  more  than  2000  feet; 
Norway  has  but  one  deep  shaft,  1900  feet;  in  South  Africa  are 
10  exceeding  1200  feet  in  depth;  and  in  the  United  States  are 
dozens  over  2000  feet  deep.  Of  these,  the  deepest  shafts  in  the 
several  nations  are  indicated  below: 

Adalbert  Przibram,  Bohemia,  Austria-Hungary 3672 

Maria  Przibram,  Bohemia,  Austria-Hungary 3360 

Produits  Colliery,  Mons,  Belgium 3937 

Viviers  Shaft,  Gilly,  Belgium 375° 

Montchanin  Colliery,  Le  Creuzot,  France 2300 


506  MANUAL  OF  MINING. 

Kaiser  William  II,  Clausthal,  Harz,  Germany 2900 

Einigkeit,  Lagau,  Saxony,  Germany 2620 

Pendleton,  Manchester,  Great  Britain 3474 

Ashton  Moss,  Manchester,  Great  Britain 3360 

Robinson  Deep,  South  Africa 1991 

Red  Jacket,  Calumet  and  Hecla,  Lake  Superior,  U.  S.  A. . .  4900 

Tamarack,  Lake  Superior,  U.  S.  A.  .  . 4450 

Lansell's  Bendigo,  Victoria 3302 

Shafts  sunk  to  facilitate  the  execution  of  long  tunnels  are 
best  located  with  their  axes  in  the  p'ane  of  the  tunnel,  afford- 
ing better  alignment,  and  only  because  of  the  difficulty  of  sup- 
porting the  shafts  at  the  tunnel  level  is  it  the  common  practice 
to  place  them  at  the  side.  Shafts  are,  however,  losing  their 
importance  for  this  work,  since  the  introduction  of  the  rapid, 
ventilating,  drilling-machines. 

As  regards  form,  the  rectangular  is  the  most  common  (Fig. 
173).  Its  timbering  is  easily  accomplished,  and  the  best  adapted 
to  loose  ground.  Where  brick  or  stone  is  used  instead  of  wood 
for  lining,  the  sides  are  arched  to  give  great  strength,  and  this 
perhaps  led  to  the  round  or  elliptical  shapes,  which  are  such 
favorites  in  Europe  on  account  of  their  greater  resistance,  and 
particularly  because  of  the  loose  soils  and  watery  strata  encoun- 
tered. That  their  entire  area  cannot  be  utilized  is,  however, 
an  objection  (Fig.  172).  The  timbering  of  the  polygonal  (12  to 
16  sides),  used  in  Belgium  and  the  north  of  France,  is  not  so 
easy  to  fit  as  is  that  in  the  hexagonal  or  octagonal  shafts. 

The  dimensions  of  the  shafts,  governed  by  the  number  of 
compartments,  should  be  carefully  studied  to  meet  all  require- 
ments of  strength,  output  and  escapement  for  a  prolonged 
period.  The  size  increases  as  the  depth  and  output  increase. 
Outputs  of  ico  tons  were  regarded  as  large  not  so  long  ago; 
but  now  many  hundreds  of  shafts  have  a  capacity  of  1000  tons 
daily.  Colliery  shafts  are  built  of  a  greater  area  than  those  in 
metal-mines,  which  latter  have  less  traffic,  besides  being  restricted 
generally  by  the  distance  between  the  walls.  The  size  of  the 


SHAFTS. 


507 


compartment  is  determined  by  that  of  the  bucket,  skip,  or  cage,, 
its  length  being  the  width  of  the  shaft,  the  length  of  which  is 
governed  by  the  number  of  divisions  (see  pages  23  and  203), 


FIG.  173. 


FIG.  174. 


FIG.  172.— Divisions  of  a  Circular  Shaft. 

Compartments  placed  side  by  side  make  a  stronger  shape  than 

if  arranged  in  a  more  compact  form  (Figs.  173  and  174).    The 

compartments    for   metalliferous 

cars  are  about  4'  X  5' ;     those  for 

the    coal-cars,  from  6  feet  to  8 

feet  wide,  by  from  10  feet  to  12 

feet  long,  measured  inside  of  the 

timbers.     The  common  sizes  for 

COal-shaftS  are    IC/X38',  and   12  The  F™r-compartment  Shaft. 

X  24',  with  wall-plates  of  some  even  50  feet  long.  In  the  Lake 
Superior  iron  region  the  shaft  dimensions  are  about  9  feet  long 
and  20  feet  wide.  In  Montana  and  Nevada  smaller  sizes  prevail, 
while  in  Colorado  a  single  compartment  suffices  for  the  small  out- 
puts of  high-grade  mineral.  The  largest  shaft  yet  begun  is  a  nine- 
compartment  shaft  38^42'  in  the  clear.  Circular  shafts  for 
buckets  holding  about  1500  Ibs.  are  8  feet  in  diameter;  for  cages 
13  feet.  The  sizes  of  the  ventilating  shafts  are  a  matter  of  in- 


508 


MANUAL  OF  MINING. 


SHAFTS.  509 

difference,  so  that  they  transmit  the  necessary  volume  of  air  with 
the  minimum  resistance,  and  at  a  current  velocity  not  exceeding 
1000  feet  per  minute.  The  upcast  shaft  is  therefore  usually  round, 
and  the  downcast  a  walled  rectangular.  Neither  should  be 
housed,  though  the  former  for  a  furnace  ventilator  may  be  pro- 
vided with  a  chimney  high  enough  to  prevent  the  distraction  of 
the  current  by  surrounding  buildings;  or  with  traps  closing 
tightly  and  quickly  if  a  fan  is  used.  An  area  of  i  sq.  ft.  for  every 
eight  men  employed  is  a  good  basis  for  the  upcast  of  a  moderate- 
sized  mine. 

The  features  governing  the  selection  of  a  site  have  already 
been  examined  on  page  20 ;  so  there  remains  to  consider  the  proc- 
ess of  sinking.  In  a  soft-ore  lode  the  shaft  section  should  reach 
from  wall  to  wall,  and  massive  shaft  pillars  be  maintained,  else 
it  is  sure  to  succumb.  In  hard-rock  lodes  the  shaft  should  prefer- 
ably be  on  the  foot-wall;'  on  the  hanging- wall  heavy  supports 
are  necessary,  especially  if  the  country  rock  is  porphyry. 

Sinking  a  Shaft. — This  process  is  slow  because  of  the  diffi- 
culty of  putting  long  angling  shot-holes.  Small  shafts  are  sunk 
by  hand  cheaper  than  by  power-drills,  and  almost  as  expeditiously, 
unless  perhaps  the  continuous  system  (see  Fig.  286)  is  used;  and 
the  loss  of  time  in  removing  all  the  implements  for  each  shot 
bears  a  large  ratio  to  the  total.  Even  in  drifting,  the  actual 
drilling  heat  is  not  more  than  half  of  the  whole  time.  The 
number  of  men  depends  upon  the  size  of  the  shaft  opened; 
only  two  miners  can  drill  to  advantage  on  an  area  of  20  sq.  ft. 
A  larger  size  gives  more  room  proportionately  to  each  miner,  and 
permits  faster  work,  and  in  a  shaft  ic/Xn'  there  is  room  for 
three  pairs  of  miners.  This  space  will  accommodate  two  machine- 
drills,  which  in  ordinary  rock  can  make  5  feet  of  advance  per 
day  (divided  into  three  shifts  of  eight  hours  each).  A  shaft 
long  in  proportion  to  its  width,  sunk  by  two  or  four  machines, 
has  two  centre-cut  ranges  of  holes  (Fig.  309),  which  are  inde- 
pendently fired.  The  cost  of  sinking  is  from  $5  to  $18  per  cubic 
yard.  Below  100  feet  the  rate  increases  each  100  feet  almost  as 
the  square  root  of  the  depth.  Rziha  says  that  in  Europe  the 


510  MANUAL  OF  MINING. 

cost  of  excavating  shafts  is  from  50  to  100  per  cent  higher  in 
wages,  and  the  cost  of  putting  in  timber  15  to  30  per  cent  higher 
in  wages,  than  the  estimate  for  the  same  amount  of  tunnel-work. 
In  the  Lake  Superior  region  one  lineal  foot  of  average  shaft 
costs  as  much  as  a  lineal  yard  of  gangway  and  a  cubic  fathom 
(216  cu.  ft.)  of  stoping. 

No  figures  can  be  given  for  calculating  the  cost  of  any  kind 
of  rockwork.  Local  conditions  of  labor  and  supplies  vary  too 
much. 

A  temporary  frame  is  erected  just  inside  of  the  permanent 
posts  of  the  head-gear  and  over  the  end  of  the  shaft  opposite 
the  pumpway.  Here  the  hoisting  machinery  is  located  and 
the  conveniences  for  loading  of  buckets.  The  pumpway  is 
partitioned  air-tight,  to  serve  for  a  ventilating- way  as  well  as  to 
contain  the  various  pipes.  Precautions  are  taken  to  prevent  the 
material  falling  down  the  shaft  by  providing  a  fencing  around 
the  opening  and  spanning  it  by  a  wide-gauge  track  upon  which 
a  platform  car  travels  to  receive  the  loaded  bucket  and  exchange 
it  for  an  empty  one.  Often  a  hood  is  provided  for  the  protection 
of  the  miners  against  fallen  rocks;  also  trap-doors  at  the  surface,, 
unless  the  ventilation  is  very  poor. 

The  work  of  excavation  is  begun  within  the  frame,  which 
latter  serves  as  the  template  for  the  remainder  of  the  timber- 
ing. At  6  or  8  feet  down,  a  temporary  scaffold  is  erected  across 
one  end  of  the  shaft  to  receive  the  material  thrown  upon  it,  whence 
it  is  lifted  to  the  surface.  When  it  has  progressed  15  feet  below 
the  surface,  the  first  permanent  curbing  is  put  into  place.  Through 
the  first  score  of  feet  the  progress  is  quite  rapid.  Beyond  this, 
progress  depends  on  the  windlass  or  engine.  For  90  feet  a 
windlass  will  suffice,  but  deeper  than  this  an  engine  must  ulti- 
mately be  used.  Except  in  small  operations  the  engine  is  placed 
at  the  start.  The  entire  bottom  is  attacked  at  once,  a  small 
corner  sump  being  carried  in  advance  for  drainage  and  for 
"bearing  in"  while  shooting. 

Extending  Shafts  by  Pentice. — Shafts  are  prolonged  with- 
out interference  to  the  regular  mining  operations,  and  without 


SHAFTS. 


danger  to  the  shaftmen,  by  opening  only  that  portion  of  the  shaft 
area  not  under  the  hoist  way  for  a  distance  of  12  or  15  feet,  and 
then  widening  it  out  to  the  entire  size  of  the  main  shaft.  This 


FIG.  176.— A  Pentice. 


leaves  a  roof  of  rock  ("pentice")  (Fig.  176),  that  shields  the 
men.  When  another  lift  has  been  started  the.  pentice  is  cut 
away,  and  another  started  for  the  next  drop. 


512  MANUAL   OF  MINING 

The  Service  of  Shaft-timbers. — There  is  no  safety  nor  econ- 
omy in  the  practice  of  leaving  a  shaft  untimbered,  even  if  the 
two  walls  are  hard  and  self-sustaining  and  can  be  dressed 
smooth.  There  is  a  thrust  from  the  walls  in  the  country  rock 
which  results  in  the  release  of  fragments  that  may  injure  the 
men  by  their  fall;  a  tendency  to  movement  exists  which  may 
completely  close  the  openings,  and  the  vibration  produced  by 
rapid  hoisting  tends  to  loosen  material;  hence  timbering  will  be 
required,  to  furnish  a  close  lining  to  the  shaft  and  a  rigid  support 
to  the  cage-guides.  The  timbers,  therefore,  are  to  serve  mainly 
as  a  means  of  preventing  movement  in  the  walls,  which,  having 
once  begun,  can  be  checked. 

The  Character  of  the  Timbering  of  Shafts,  whether  vertical 
or  inclined,  depends  upon  the  speed  desired  for  hoisting.  It  is 
more  substantial  for  cage-hoisting,  or  mechanical  haulage,  than 
it  would  be  for  a  bucket-hoist  or  animal  haulage.  The  varying 
conditions  as  to  the  character  of  the  rock,  the  area  of  the  shaft, 
and  the  firmness  of  the  enclosing  strata,  determine  the  choice  of 
system  to  be  employed.  An  opening,  for  the  shaft  made  any 
greater  than  is  necessary  to  accomplish  the  desired  end,  not 
only  increases  the  expense  of  opening,  but  also  the  cost  of  sub- 
sequent maintenance  The  timbers  become  needlessly  long; 
the  cost  of  breaking,  high;  and  the  material  to  be  transported, 
larger  than  necessary.  Moreover,  the  cost  of  back-filling  the 
useless  space  would  be  materially  increased  and  the  size  of  the 
timbers  in  a  horizontal  direction  must  be  proportionally  larger 
for  a  given  rate  of  pressure. 

Timbering  a  Shaft.  —  The  timbering  of  a  shaft  may  be 
done  simultaneously  with  or  subsequent  to  the  process  of  sinking, 
according  to  the  firmness  of  the  ground  through  which  the  work 
is  carried.  Through  loose  ground,  or  for  the  first  30  or  40  feet 
of  depth  through  the  alluvium,  the  timbering  is  placed  as  promptly 
as  possible  and  of  a  thickness,  compactness,  and  strength  de- 
pending upon  the  nature  of  the  ground,  whether  loose  and  rocky 
or  wet  and  sandy.  This  continues  until  hard  rock  is  reached,, 
after  which  some  form  of  timber-frame  is  introduced.  If  any 


SHAFTS. 


water  is  encountered,  water-rings  or  coffer-dams  are  built  after 
the  usual  plan  (Fig.  177)      At  the  line  of  union  between  the 


FIG.  177. — A  Clay -puddled  Shaft-crib. 


FIG.  178. — The  Timber-frame  of  a  Shaft. 

alluvium  and  the  bed-rock  a  water-tight  joint  is  made.  The 
shelf  of  rock  is  dressed  level  at  a  sufficient  distance  below  the 
upper  surface  of  the  stratum  to  allow  for  the  packing,  and  on  it 


514  MANUAL  OF  MINING. 

is  built  a  wedging  curb  of  segments  of  heavy  timbers.     This 
curb  is  tightly  wedged  and  packed  with  puddled  clay. 

The  standard  frame  consists  of  four  pieces  (Fig.  178),  each 
timber  of  the  set  being  known  as  the  plate.  At  the  sides,  RR, 
are  wall-plates,  and  the  short  plates  at  the  ends  are  end-plates. 
If  the  shaft  is  to  be  divided  into  compartments  it  is  done  by  the 
use  of  buntons,  BB,  or  girts  bolted  to  or  gained  into  the  wall- 
plate.  The  greater  the  length  of  the  wall-plates  the  stouter  are 


FIG.  179. — Boxed  Joint.  FIG.  180. — Boxed  Shoulder. 


FIG.  181. — The  Boxed-shoulder  Crib. 

the  buntons  or  girts.  On  the  inside  of  the  end-plates  and  the 
girts  are  spiked  the  guides  for  the  hoisting-cages.  If  any  com- 
partment is  to  serve  as  the  travelling-  or  the  pumping-way,  or 
even  for  ventilation,  it  is  partitioned  off  by  planks  nearly  vertical 
to  branches,  as  in  ihe  right-hand  compartment  (Fig.  175). 

Timber  Joints. — The  joints  in  the  frames  are  of  several  forms, 
regard  being  had  to  the  expense  of  framing  and  the  convenience  of 
handling  the  pieces  as  well  as  the  efficiency  of  the  joint  for  resist- 
ing pressure  The  timbers  may  be  rough-hewn,  particularly  if 
solid  cribbing  work  is  to  be  erected,  but  preferably  with  their 
lengths  dressed  to  a  particular  template.  In  firm  non-decom- 


SHAFTS.  515 

posing  ground  the  timbers  experience  little  pressure,  and  there- 
fore stability  rather  than  strength  is  sought,  the  latter  being 
secured  by  ample  shaft  pillars.  Under  such  circumstances  a 
line  of  stiff  guide-planks  will  be  sufficient,  placed  in  close  con- 
tact, or  in  a  frame,  according  to  the  liability  of  the  material 
from  walls  of  the  rock  to  scale  off  and  fall  into  the  shaft  way. 

The  simplest  form  of  joint  for  such  a  lining  consists  of  2-inch 
planks  cut  to  a  square  butt-joint  at  the  ends  to  a  correct  length 
and  placed  with  their  longer  dimensions  vertical,  so  that  the  alter- 
nating sets  break  joints  (Fig.  179);  in  other  words,  the  wall- 
plates  of  one  set  overlap  and  rest  upon  the  end-plates  of  the  set 
below,  and  in  turn  support  the  ends  of  the  end-plates  of  the  set 
above.  The  end-plates  are  correspondingly  cut.  It  is  efficient 
and  cheap  for  curbing  shallow  shafts.  It  is  finished  by  spiking 
firmly  from  top  to  bottom  four  triangular  corner-pieces. 

Shaft-linings. — In  Figs.  180  and  181  is  illustrated  a  form  of 
curbing  adopted  for  somewhat  heavier  pressure,  in  which  the 
four  pieces  in  the  set  are  cut  to  the  square-box  joint  to  ex- 
actly match.  These  timbers  are  4  inches  wide  by  6  or  8  inches 
high.  In  Fig.  182  is  illustrated  also  a  somewhat  similar  system 
for  slightly  heavier  timbers,  in  which  the  framing  of  the  boxed 
shoulders  of  the  four  pieces  is  vertical  instead  of  horizontal. 
If  these  are  cut  to  template,  it  is  not  necessary  that  the  timbers 
should  match  in  height  for  each  set.  Corner-pieces  are  used 
inside  of  this  frame. 

This  form  of  casing  is  built  in  sections  of  30  feet  each  in 
height  by  placing  a  pair  of  lo-inch  stulls  in  a  securely  horizontal 
position  and  building  up  from  them  the  sets  as  indicated. 
During  the  progress  of  this  timbering  the  corners  are  plumbed 
continually,  and  waste  rock  is  packed  closely  with  its  upward 
progress.  Two  men  can  complete  one  section  of  a  shaft  5  feet  by 
9  feet  in  dimensions  in  four  days'  time  with  one  or  two  helpers 
at  the  packing.  The  men  are  supported  on  a  cradle  suspended 
by  a  rope  from  the  upper  stulls. 

Two- compartment  shafts  can  also  be  lined  by  this  form  of 
casing,  when  buckets  are  used  for  hoisting,  but  for  cage  use  the 


MANUAL  OF  MINING. 


timbers  are  not  secure.  Cribbing  is  used  for  small  shafts,  as  it 
is  cheap  and  simple,  but  would  be  too  expensive  and  cumber- 
some in  shafts  of  two  or  three  compartments. 

It  is  not  necessary  <hat  these  timbers  should  be  in  close  con- 
tact, but  the  framing  may  be  made  such  that  a  slight  open  space 
is  left  between  each  set  of  timbers  by  leaving  the  tenons  larger 
than  is  necessary  for  the  shoulders  on  the  ends. 

A  more  efficient  joint  is  shown  in  Fig.  183,  when  the  pressure 
from  the  wall-rock  is  great  enough  to  require  the  lining  of  con- 


FlG.  182. — The  Halved  Joint. 


FIG.  183. — The  Bevel  Hitch. 


FIG.  184.— The  Boxed  Joint. 


siderable  thickness.  In  this  casing  a  half-shoulder  is  cut  at  each 
end  to  a  true  fit,  and  the  pieces  laid  for  smaller  timbers.  The 
bevel  hitch  at  the  ends  weakens  the  timber  unless  its  full  strength 
be  obtained  by  bringing  it  to  a  close  bearing  with  the  bevel 
face  of  its  mate. 

This  character  of  casing  may  be  employed  when  passing 
through  dry  quicksand  or  running  strata,  in  which  case  the 
planks  would  be  spiked  upward  from  below  as  fast  as  they 
were  inserted.  This  curbing  is  arranged  with  the  layers  in 


SHAFTS. 


5*7 


alternate  thicknesses,  measured  in  a  horizontal  direction,  so 
as  to  present  the  uneven  surface  at  the  back  of  the  lining 
against  and  upon  which  the  sand  may  press.  For  example,  the 


FIG.  185. — Shaft-timbering  at  a 
Landing. 


FIG.  186.— Solid  Shaft-lining  with 
Bevel  Hitch. 


planks  may  be  in  sets  2  inches  high  and  8  inches  thick  all  around, 
alternating  with  planks  2  inches  high  and  TO  inches  thick. 


5l8  MANUAL   OF  MINING. 

In  exceptional  cases  this  form  of  timbering  may  stand  safely 
without  corner-pieces,  but  it  is  wiser  to  employ  them,  utilizing  also 
waste  packing  to  keep  the  joints  in  position.  This  may  be  also 
arranged  in  such  a  manner  as  to  leave  intervals  between  the 
sets  through  which  some  material  can  be  removed  to  relieve 
the  unusual  pressure. 

Another  method  of  open  framing  of  wall-plates  resting  in 
the  end-pieces,  much  like  the  log-cabin  style,  but  without  any 
notches  at  the  ends  of  the  pieces  is  in  frequent  use  in  firm 
ground. 

Where  the  strata  are  self-supporting,  the  planks  need  not  be 
more  than  4  inches  thick,  set  on  the  edge  for  a  depth  of  150  feet, 
or  3  inches  thick  for  a  depth  of  120  feet;  but  if  the  ground  is 
soft,  or  crumbles,  or  there  is  a  flow  of  water,  other  provisions 
must  be  made  to  increase  the  strength  of  the  set  or  to  provide  a 
supplementary  dam  to  relieve  it. 

It  is  difficult  to  determine  the  requisite  size  of  the  timbers 
in  a  shaft,  for  the  amount  of  pressure  from  the  walls  cannot  be 
measured.  Experience  with  the  character  of  the  rock,  its  strength 
and  freedom  from  cleavage  planes,  can  alone  furnish  any  guide 
to  the  engineer.  The  dimensions  of  the  timbers  are  there- 
fore fixed  by  other  conditions,  allowance  being  made  for  emer- 
gencies. 

Square-set  Timbering. — Shafts  are  more  commonly  framed 
in  square  sets  (Fig.  187).  Square-set  timbers  for  shafts  are  8 
inches  square  and  consist  of  side  wall-plates,  end-plates,  and 
four  posts  supporting  the  horizontal  frame.  A  set  consists  of 
four  plates  and  as  many  girts,  buntons,  or  inner  struts  as  there 
-are  hoist  compartments.  A  three -compartment  shaft  is 
shown  in  Fig.  188  with  lagging  for  lining.  The  plate-joints 
are  halved  with  a  hitch,  or  square  shoulder-notch,  i  inch  deep, 
•cut  into  the  tenon  for  a  support  to  the  post.  A  centre  line  is 
marked  on  the  inside  face  of  the  plates,  from  which  all  measure- 
ments are  taken  for  tenons,  mortises,  and  mitres,  as  well  as  for 
plumbing. 

The  timbers  are  prepared  by  boring  four  i-inch  holes,  by  a  tern- 


SHAFTS. 


U[ 


id 


n 


n 


g 
3 

n 


Qsic 


FlG.  187. — A  Four-compartment  Square  Shaft. 


520 


MANUAL   OF  MINING. 


plate,  near  the  ends  of  the  wall-plates.  Through  these  are  forced 
iron  rods  with  a  long  screw-thread  at  one  end  and  a  hook  at  the 
other.  The  length  of  these  rods  is  something  more  than  the 
half  distance,  out  to  out  vertically,  of  the  frames.  Some  of 


FIG.  188. — A  Lagged  Lining. 

the  rods  have  a  ring  at  one  end  instead  of  a  hook,  and  other 
rods  in  pairs  are  made  of  a  length  of  the  odd  multiples  of  the 
half  distance.  A  set  of  liberal-sized  washers  and  nuts  is  sup- 
plied for  each  pair  of  wall-plates. 

Setting  the  Timbers. — If  the  shaft  is  started  on  the  hillside, 
affording  height  for  the  convenient  disposal  of  the  debris,  a 
supporting  frame,  from  which  is  to  be  suspended  the  first  section 
of  timbers,  is  laid  on  the  surface.  If  started  on  level  ground, 
where  height  must  be  obtained  artificially,  a  top  collar  is  laid 
on  the  timbering  above  the  surface,  which  timbers  are  heavy, 
and  in  a  frame  of  four  pieces  with  the  shaft -opening  in  the  middle. 
This  frame  is  then  reinforced  by  an  outer  crib,  between  which  is 
a  filling  of  rock.  The  whole  structure  must  be  blocked  firmly. 
If  the  ground  is  loose,  the  side -plates  are  extended  far  beyond 


SHAFTS.  521 

me  limits  of  the  shaft  and  bolted  to  heavy  cross-sills  for  support 
till  the  firm  ground  has  been  reached. 

The  upper  set  having  been  placed  level  with  two  of  the  sus- 
pension-rods in  each  wall-plate  with  hooks  down,  two  wall- 
plates  of  the  next  lower  set  are  attached  with  their  hooks  up. 
The  end-plates  are  then  laid  on  the  wall-plates,  the  posts  are 
cct  in  their  hitches,  and  the  nuts  are  screwed  down  tightly. 
The  posts  are  plumbed,  blocks  inserted,  and  wedges  driven. 
Spiked  to  the  back  of  each  plate  is  a  2"  X4"  strip,  on  which 
rests  the  lagging  or  lining.  After  the  filling  has  been  stowed  the 
process  is  repeated  for  the  next  lower  set. 

Two-compartment  shafts  are  made  in  two  divisions  of  the 
same  size,  framed  identically  alike,  the  longer  wall-plates  being 
strengthened  by  girts  and  centre-posts. 

In  the  three-compartment  shafts  the  one  assigned  to  pump- 
ing is  larger  than  those  intended  for  hoisting,  and  its  girts  and 
posts  are  more  rigid  than  at  the  other  end.  The  process  is  the 
same  as  for  the  single-compartment,  the  buntons  being  fitted 
into  the  mortises  in  the  wall-plates  after  the  wall  and  end-pieces 
have  been  laid.  The  wall-plates  should  be  in  one  piece  unless 
inconvenient.  But  if  in  two  pieces,  the  short  one  should  be 
at  the  pumpway,  with  the  divisional  girt  located  at  the  joint. 
The  eight  posts  are  placed,  bound,  blocked,  wedged,  and  lined  up. 

The  four-cornpartment  shafts  with  three  hoistways  are  in  single 
line,  the  fourth  being  used  for  lowering  timbers  and  machinery, 
and  the  framing  is  identical  with  that  for  three  compartments. 
Those  constructed  with  the  compartments  in  two  rows  (Fig.  174) 
have  their  errd-plates  longer  to  enclose  a  double  width  (Fig.  187). 
In  this  event  the  stations  are  in  the  corners  and  the  direction  of 
the  shaft-frame  is  determined  by  the  subterranean  requirements. 
The  excavation  is  square,  the  timbering  compact  and  rigid  and 
not  excessively  long,  the  shaft  has  greater  capacity  as  compared 
with  the  straight  line,  and  there  is  but  5  cu.  ft.  more  of  ground 
removed  per  foot  of  shaft  than  in  the  latter.  The  middle  fram- 
ing in  the  direction  of  the  hoist  is  often  braced  with  diagonal 
struts  to  solidify  the  sets. 


5 2-'  MANUAL  OF  MINING. 

Timbering  in  Firm  Ground.— If  the  rock  is  firm  enough  to 
admit  of  sinking  50  feet  or  so  before  commencing  the  timber- 
ing, the  process  can  be  more  cheaply  conducted  by  building 
upward  from  reachers  or  stulls  set  into  both  walls.  On  these, 
four  pieces  are  framed  to  the  studdles  or  struts  at  the  corners 
and  at  the  compartment  partitions.  On  these  struts  a  similar 
set  is  framed  6  feet  above,  to  support  another  parallelepiped, 
and  so  on  up.  Planks  ("lagging")  are  driven  in  around  these 
frames,  and  the  spaces  to  the  rock  filled  with  broken  waste.  The 
joints  of  each  timber  are  of  the  pattern  shown  in  Figs.  189  and 

FIG.  189.  FIG.  190. 


I 


c 


Joints  of  Square-set  Timbers. 

190.  The  end-plates  and  struts  are  usually  8  inches  square, 
while  the  wall-plates  are  laid  8  inches  vertically  and  10  or  12 
inches  horizontally.  Fig.  177  illustrates  another  form  of  tim- 
bering rectangular  shafts  with  vertical  corner-plates  and  hori- 
zontal lagging. 

Shafts  such  as  the  Comstock,  6^X24',  for  continuous  heavy 
hoisting,  are  fitted  with  timber  as  much  as  14  inches  squaoe,  and 
lagged  with  3-inch  plank.  Where  friable  rock  is  penetrated, 
the  frames  are  braced  by  inclined  struts  that  prevent  settlement. 
When  the  ground  is  friable,  marly,  or  wet,  the  methods  assume 
a  caisson  character.  Another  plan  comprises  a  stout  framing 
(C,  Fig.  177),  inside  of  which  is  another  strong  planked  cribbing, 
between  which  clay  is  puddled  to  exclude  the  water.  The  B.  &  O. 
shaft  at  Taylorsville,  Ind.,  was  thus  successfully  carried  through 
quicksand;  the  outside  crib  was  of  12-inch,  the  inside  of  ic-inch 
timbers,  with  a  4-inch  puddled  wall.  The  famous  Hollenback 
shaft,45/4"Xn'6"  inside,  has  a  1 2-inch  clay  wall  for  31  feet  of 
depth  (Fig.  191).  It  was  designed  for  a  daily  output  of  2500  tons 
of  coal. 


SHAFTS. 


S23 


$24  MANUAL  OF  MINING. 

Repairing  Timbers.  —  If  the  timber  shows  signs  of  giving 
way,  other  means  of  securing  the  shaft  must  be  invoked,  \\ith 
expert  timbermen  the  joints  may  be  strengthened  or  the  frame 
replaced,  but  it  is  preferable  to  reinforce  them  by  intermediate  sets. 

After  supporting  the  several  sets  on  either  side  of  those  to  be 
repaired,  the  faulty  pieces  can  be  removed  one  by  one,  excavating 
enough  ground  from  behind  each  post  to  allow  of  its  being  driven 
back  from  the  shaft  clear  of  the  timbers  and  then  replacing  each 
piece  singly,  wedging  and  driving  it  into  place  from  behind.  It 
may  be  possible  also  to  chop  away  the  piece  if  it  is  not  desirable 
to  open  up  more  ground  behind  the  set.  The  new  post  is  wedged 
into  position  before  the  next  post  is  inserted.  When  plates  are 
to  be  removed  and  new  ones  inserted,  the  posts,  the  lagging,  and 
the  blocks  supporting  that  side  must  be  removed  and  one  piece 
inserted  at  a  time.  If  it  bends  with  the  ground  and  will  not 
stand  during  the  pressure,  false  timbers  must  be  used  to  hold 
the  walls  in  place. 

Building  Landing-stations. — The  stations  are  prepared  at 
various  depths  of  about  100  feet  each,  from  which  the  levels  are 
run  into  inner  workings  of  the  mine.  These  stations  must  be  in 
excavations  large  enough  to  serve  as  centres  of  distribution  of 
cars  into  the  rooms  and  of  storage  of  the  loaded  cars  preliminary 
to  hoisting.  Their  height  is  necessarily  greater  than  in  the  level 
and  their  width  is  at  least  that  of  the  longer  dimension  of  the 
shaft.  The  timbering  and  lining  of  such  working  stations  is  a 
component  part  of  the  shaft  system,  with  such  modifications  as 
are  necessary  to  connect  it  with  the  balance  of  the  works,  besides 
affording  a  full,  clear,  unobstructed  width  for  free  movement  of 
cars  and  men.  The  timbers  of  the  stations  usually  consist  of 
the  four-piece  level  set,  the  caps,  however,  being  very  heavy, 
frequently  of  iron.  The  lagging  at  the  roof  is  of  3-inch  plank. 
A  height  equal  to  that  of  two  sets  of  shaft-timbers  will  afford 
ample  height  for  handling  timbers,  pump-pipe  lengths,  etc. 

In  opening  a  station,  four-hitch  timbers  are  wedged  under 
the  last  set  of  the  shaft  crosswise,  and  the  shaft  continued  some 
distance  below.  A  three-post  cap  and  sill-set  is  placed  against 


SHAFTS.  525 

the  shaft-timbers  and  finished.  Portions  of  the  wall-plates 
lying  between  the  shaft  and  the  station,  breast  high,  are  then 
sawed  out,  after  which  the  timbers  may  be  continued  backward, 
slanting  downward  as  far  as  required  to  usual  dimensions  of  the 
drift  (Fig.  185). 

For  inclines  the  stations  are  of  two  kinds:  one  arranged  so 
that  the  ore-cars  will  dump  directly  into  the  hoisting-skip;  the 
other  is  a  large  ore-bin  built  below  the  station  at  one  side,  from 
which  a  discharge  can  be  made  into  the  skip  at  proper  intervals. 
The  latter  is  far  more  economical  in  the  long  run.  The  skip- 
loading,  moreover,  is  independent  of  the  haulage  system.  But 
the  excavation  must  be  larger  than  for  the  former. 

The  sets  are  lined  up  by  straight-edge  and  plumb-bobs,  the 


FIG.  192. — Station-cribbing  on  a  Slope. 

former  being  long  enough  to  overreach  the  distance  between  two 
sets.  A  frame  can  be  built  in  such  a  manner  that  a  plumb-bob 
will  swing  free  within  small  limits  from  the  upper  end  of  the 
straight-edge,  the  lower  member  resting  upon  the  bottom  sill. 

In  Fig.  192  is  shown  a  station  built  in  cribwork. 

If  the  expense  would  warrant  it  and  the  diminished  area  is 
not  objected  to,  a  second  lining  might  be  inserted  and  prove 
secure.  In  the  Lake  Superior  region  mine  caissons  were  in^ 


526  MANUAL  OF  MINING. 

voked  after  various  futile  experiments  had  been  made  with  other 
methods  of  reinforcing  the  shaft-timbers.  For  example,  iron 
caissons  were  forced  down  inside,  or  outside,  of  the  old  timbers, 
according  as  the  ground  was  stiff  or  soft.  In  the  latter  it  was 
even  possible  to  drive  them  by  their  own  weight,  with  very  little 
supplementary  power.  These  cylinders,  in  segments  and  sec- 
tions, were  added  one  by  one  at  the  surface,  keeping  pace  with 
the  progress  of  sinking.  In  one  instance  a  cast-iron  cylinder 
15  feet  in  diameter  and  an  inch  thick  was  forced  down  at  a  rate 
of  2  feet  per  hour  through  80  feet  of  morainal  matter 

Forepoling. — In  very  loose  ground  the  timbering  must  be 
conducted  promptly  by  some  method  of  shaft-lining.  The  usual 
plan  is  that  of  spiling  (Fig.  193).  A  set  of  planks  is  driven  nearly 
vertically,  blocked  and  wedged  into  position.  Each  piece  bears 
against  the  walls  and  is  fixed  as  the  portion  of  a  complete  lining  to 
the  entire  shaft.  In  soft  or  running  ground  which  cannot  be  tim- 


FIG.  193. — The  Forepoling  of  a  Shaft. 

bered  or  requires  timbers  excessively  heavy  for  lining,  the  method 
consists  in  holding  back  the  dangerous  ground  below  the  last 
set  in  position  by  an  enclosing  and  protecting  shell  of  forepoling, 
as  described  in  Chapter  III  for  levels  (Fig.  271).  The  forepole  is 
a  stout  plank  sharpened  at  its  foot,  often  shod  with  iron,  and  has 
a  length  greater  than  that  of  the  distance  across  two  sets.  It  is 
driven  by  a  sledge,  a  single  plank  at  a  time,  from  the  back  of.  the 
plates  in  the  last  and  lowest  set.  Little  by  little  the  material 
is  removed  to  allow  of  it  being  driven  a  short  distance.  Each 


SHAFTS. 


527 


plank  around  the  shaft  is  driven  in  turn  until  the  entire  shell 
has  been  advanced.  This  process  is  repeated  until  a  new  set 
can  be  inserted,  behind  which  and  in  front  of  the  previous  lining 
of  four  poles  is  driven  another  set  in  a  similar  manner.  The 
forepoling  is  started  at  a  considerable  angle  outward,  but  in  its 
downward  progress  is  forced  toward  the  vertical  until,  by  the 
time  an  advance  has  been  made  for  another  set  to  be  placed, 
the  planks  are  nearly  in  their  desired  position,  giving  security 
to  the  frame  and  affording  safety  to  the  men.  When  the  second 
set  of  spiles  has  been  started  those  constituting  the  first  set  are 
driven  their  full  length. 

Forepoling  has  proven  very  successful,  though  it  requires  con- 
siderable timbering.     It  is  a  perfectly  safe  method  and  can  be 


FIG.  194. — Forepoling. 

employed  in  almost  any  variety  of  dangerous  ground  except 
quicksand.  Frequently,  when  the  ground  proves  treacherous, 
there  is  perhaps  too  rapid  a  rise  of  the  bottom  of  the  shaft.  In 
this  event  the  simplest  plan  is  to  brace  the  bottom  sets  and  floor 
the  foot  of  the  shaft  over  its  entire  area,  leaving  only  a  small 
opening,  advanced  first  by  forepoling  (Fig.  194),  and  subse- 


528  MANUAL  OF  MINING. 

quently  enlarged  to  the  full  dimensions  of  the  shaft.  This  plan 
resembles  somewhat  that  of  the  Anderson  pilot  tube,  as  illustrated 
in  Chapter  IV. 

Wood  Tubbing  for  Circular  Shafts. — For  circular  shafts  the 
false  work  descends  with  the  shaft,  but  the  cribbing  is  built 
upward  in  sections  from  stout  reachcrs  of  timber,  bedded  when- 
ever suitable  foundation  offers,  or  from  a  properly  dressed  ledge 
of  the  rock,  and  firmly  wedged  against  the  sides.  The  timbers, 
assuming  the  character  of  voussoirs,  hooped  with  iron,  may  be 
mere  short  blocks  or  wedges,  or  they  may  be  longer  timbers  form- 
ing a  polygon  with  less  number  of  sides.  In  the  latter  case  they 
are  united  by  iron  dogs.  In  ordinary  ground  the  sets  are  held 
apart  by  vertical  props  with  a  solid  packing  behind  them.  The 
solid  walling  may  also  be  suspended  from  a  heavy  frame  at  the 
surface  by  iron  rods  (Fig.  195).  In  any  event  the  joints  and 
fitting  receive  the  greatest  care,  and  many  of  the  old  shafts  are 
high  types  of  the  carpenter's  art. 

The  increasing  scarcity  and  cost  of  large  timbers,  the  ex- 
pense of  fitting  and  maintenance,  their  short  life,  and,  finally, 
the  corrosion  of  spikes  and  splice-plates,  with  the  consequent  leak- 
ages, have  caused  the  abandonment  of  wood  tubbing,  and  the 
adoption  of  iron  and  masonry  for  all  permanent  ways.  The 
effect  of  the  hot  atmosphere  of  the  mine  upon  timbers  is  a  decom- 
position that  is  not  always  detected  on  the  surface,  but  once 
begun,  only  better  ventilation  can  delay  ultimate  destruction. 
Dry  timbers  should  be  frequently  probed;  alternations  of  wet 
and  dry  are  exceedingly  destructive;  wet  timber  will  last  longer 
than  dry.  Preservatives  have  been  attempted,  with  much  suc- 
cess. In  salt-mines  steeping  in  brine  gives  great  endurance. 
The  sulphates  and  chlorides  of  zinc  have  proven  excellent  anti- 
septics; and  a  grand  opening  offers  to  the  discoverer  of  a  means 
of  freeing  the  lead  ores  of  the  Western  States  of  the  obnoxious 
zinc,  and  at  the  same  time  utilizing  it  as  a  preservative. 

Masonry  Walling  of  Shafts. — The  use  of  masonry  involves 
but  one  disadvantage:  it  presupposes  ground  that  will  stand 
safely  for  a  couple  of  weeks  without  much  support.  Before  the 


SHAFTS. 


529 


permanent  structure  can  be  introduced,  a  considerable  depth 
must  be  reached,  to  obtain  a  sure  foundation  upon  reachers,  or 
upon  a  ledge  from  which  the  masonry  is  erected,  the  temporary 
timbering  and  bracing  being  gradually  removed  as  the  construe- 


I      <         1 

#  {»   I       14  V 


FIG.  195. — Building  the  Shaft  Walling. 

tion  proceeds.  When  a  very  secure  ledge  or  base  cannot  be 
had,  a  wedge-shaped  chamber  is  built  for  some  distance  back 
into  the  rock  from  which  the  solid  crib  supports  the  walling. 

If  the  pressure  from  the  walls  is  not  very  great,  walls  are 
built  up  of  brick  or  rock,  packed  securely  behind  by  waste.     Often 


533  MANUAL  OF  MINING. 

the  mine-water  carries  matter  in  solution  that  cements  the  whole 
into  one  solid  mass.  When  great  pressure  is  expected,  the  sides 
are  arched  toward  it;  and  in  very  bad  ground  all  four  sides  are 
curved,  or  the  circular  form  is  adopted.  The  arc  should  be  such 
that  its  chord  is  perpendicular  to  the  direction  of  pressure.  In 
such  event,  the  foundations  for  the  sections  are  curbs  of  over- 
lapping timbers  patterned  to  the  curve,  or  of  late  years  of  cast 
iron,  with  slabs  of  wood  at  the  joints.  The  packing  behind  is 
carried  up  with  the  brick  or  masonry  until  the  ledge  of  the  upper 
section  is  reached,  when  it  is  removed  gradually  and  the  two 
sections  united.  In  some  instances  the  masonry  compartments 
are  built  at  the  surface  and  lowered  into  place.  Brick  is  well 
adapted  for  quick  arch  work.  The  wall  of  a  shaft  13  feet  in 
diameter  is  four  half-bricks  thick;  the  labor  of  laying  it  from 
a  staging  is  one  and  one-half  days  per  M.  The  masonry  is 
supported  by  rods,  b  (Fig.  195),  from  beams  a,  a,  buried  firmly 
in  the  walls. 

Masonry  is  heavy  to  support,  and  not  any  cheaper  now  than 
iron,  with  which  many  shafts  are  successfully  curbed.  Rings 
of  I  beams  or  channel-bars  form  the  curbs,  upheld  at  proper 
distances  apart,  by  struts  of  wood  or  iron,  and  backed  by  heavy 
planks  or  T3j  sheeting  (Fig.  195).  English  engineers  use  old 
railroad  iron  similarly.  Prepared  at  the  surface,  the  curbs  may 
be  lowered  into  place  and  quickly  set,  with  little  labor.  A  pack- 
ing of  concrete  is  used  at  Saarbruck,  giving  increased  strength 
and  durability.  It  is  estimated  that  the  initial  cost  of  iron  lining 
in  place  is  twice  that  of  wood  and  equal  that  of  masonry,  but 
the  cost  of  maintenance  is  one  third  that  of  wood  and  nearly 
the  same  as  with  masonry,  if  the  shaft  is  dry. 

Shaft-bottoms. — A  very  important  feature  in  shaft-mines  is 
the  arrangement  at  the  bottom  of  the  shaft  for  receiving  and 
disposing  of  the  cars.  The  loaded  and  the  empty  cars  should 
go  on  and  off  at  opposite  sides  of  the  cage,  as  on  the  surface. 
Hence  the  heading  and  the  position  of  the  engine  will  be  parallel. 
As  on  the  surface  also,  the  grades  to  the  cage  and  from  it  should 


SHAFTS.  531 

TDC  downward.  The  loading  can  then  be  done  expeditiously  and 
economically,  for  not  only  is  the  cost  of  handling  less,  but  the 
capacity  of  the  shaft  is  increased  thereby.  By  this  means  the 
work  of  caging  can  be  performed  by  fewer  men  than  if  the  tracks 
were  on  a  level  or  up  grade  for  the  empties.  The  latter  are  run 
down  to  some  convenient  point  where  they  are  picked  up  by 
the  locomotive  or  mules. 

If  it  happens  that,  in  order  to  accomplish  this,  the  shaft- 
bottom  must  be  lowered  and  the  main-entry  bottom  also  taken 
up,  the  subsequent  saving  will  warrant  the  initial  expense.  Occa- 
sionally it  is  found  desirable  to  build  a  special  roadway,  passing 
the  sides  of  the  cage,  to  return  the  empties  to  their  appropriate 
entries. 

The  headings  at  the  shaft-bottom  are  usually  arched  with 
brick,  with  considerable  packing  back  of  the  arch,  and  further 
protection  of  longitudinal  layers  of  4-inch  plank.  Care  should 
be  taken  that  the  space  between  the  rock  and  the  back  of  the 
arch  be  completely  filled  to  avoid  movement  or  side  pressure  of 
the  roof.  The  supporting  walls  to  the  springing  line  should 
not  be  over  4  feet  high,  and  the  curve  of  the  arch  as  low  as  will 
give  sufficient  height  for  the  work  to  be  done  at  the  shaft.  Not 
infrequently  an  inverted  arch  is  built  for  the  floor. 

Shaft  Pillars.  —  The  dimensions  of  the  shaft  pillars  should  be 
such  as  to  ensure  absolute  security  to  the  works.  Though  many 
local  conditions  affect  their  size,  the  following  rule  may  be  given 
for  flat  seams,  where  the  depth  does  not  exceed  1000  feet.  With 
D  to  represent  the  depth  of  the  shaft,  t  the  thickness  of  the  seam, 
and  R  the  radius  of  the  pillar,  all  dimensions  being  in  feet,  then 


In  inclined  seams  it  Is  evident  that  the  larger  portion  of  the 
pillar  should  be  to  the  rise  rather  than  to  the  dip.  Its  length 
then  is  greater  than  the  dip  portion  by  three  fourths  of  d,  in 
which  d  is  the  distance  in  feet  along  the  seam  from  the  foot  of 


532  MANUAL  OF  MINING. 

the  shaft  to  the  foot  of  a  perpendicular  drawn  from  the  top  of 
the  shaft  to  the  seam.  Then,  if  R  is  the  length  of  the  pillar  to 
the  dip, 

R  =  6o+o.o$D\/T. 

The  line  at  right  angles  to  the  dip  of  the  seam  is  the  ultimate 
breaking  line. 

In  a  seam  6  feet  thick,  with  an  inclination  of  one  in  three,  in 
a  shaft  1200  feet  deep,  the  dimensions  of  the  dip  pillar  will  be  207 
feet.  The  value  for  d  being  346,  the  length  of  the  pillar  to  the 
rise  will  be  406.5  feet.  The  total  length  of  the  pillar  becomes 
then  673.5  feet  and  its  width  for  one  shaft  twice  that  of  the  dip 
pillar,  or  414  feet;  for  two  shafts,  180  feet  apart,  594  feet. 

In  the  case  of  a  fresh  opening  in  another  seam  of  coal  above 
or  below  the  bottom  already  existing,  care  should  be  taken  that 
the  headings  in  both  seams  be  not  placed  immediately  above 
one  another  and  parallel,  unless  the  distance  is  very  great. 

The  underground  s  ables  should  be  in  a  position  as  con- 
venisnt  as  possible,  and  well  ventilated  with  splits  in  each  sepa- 
rate stable  for  24  animals.  The  archway  overhead  should  be 
12  feet  by  16  feet. 

The  following  references  are  cited: 


REFERENCES. 

Earth  Pressures  in  Deep  Mines,  Coll.  Guard.,  Jan.  2,  1903. 

The  Framing  of  Rectangular  Shaft  Sets,  Wilbur  E.  Sanders,  Eng.  &  Min. 
Jour.,  Mar.  10,  1904;  Timbering  Shaft  Coal  Mines,  Wm.  Bradford,  Coll. 
Mgr.,  Nov.  1896;  Shaft  Sinking  in  Germany,  H.  Huhn,  Coll.  Guard.,  LXXII, 
932;  Shaft  Linings,  Coll.  Guard.,  Mar.  27,  1003;  Plumbing  Deep  Shafts, 
M.  &  M.,  Vol.  XXII,  247;  Shaft  Sinking  Contracts,  M.  &  M.,  Vol.  XXIII, 
260;  Shaft  Sinkin-  Progress:  Four  Shafts  Witwatersrand,  E.  &  M.  J.,  May 
22,  1897,  507;  Shaf  Timberi-g,  Leith  Coal  Shaft,  Coll.  Eng.,  Aug.  1896,  3. 

The  Safety  Aspect  of  Inclines,  Henri  Ghysen,  Coll.  Guard.,  Mar.  30, 1000: 
Dipping  Shaft,  M.  &  M.,  Vol.  XX,  15. 

Automatic  Shaft  Doors,  H.  Siede,  Gluckauf,  Oct.  31,  1003;  Methods  of 
Closing  the  Tops  of  Upcast  Winding  Shafts,  A.  Ried,  Fed.  Inst.  M.  E.,  X; 


SHAFTS.  533 

Some  Arrangements  for  Preventing  Accidents  at  Level  Landings  in  Cage 
Dips  and  Shafts,  A.  R.  Sawyer,  N.  Staff.  Inst.,  VIII,  204. 

Deep,  Level  Shafts  on  the  Witwatersrand,  with  Remarks  on  the  Method 
of  Working  the  Greatest  Number  of  the  Deep  Level  Mines  with  the  Fewest 
Possible  Shafts,  I.  H.  Leggitt,  A.  I.  M.  E.,  Vol.  XXX,  947;  Cost  of  Mining 
and  Depth,  A.  C.  Lane,  Mineral  Industry,  IV,  777;  How  Deep  can  We 
Mine?  A.  C.  Lane,  Mineral  Industry,  IV,  767;  Deep  Shafts  of  the  World, 
B.  H.  Brough,  Coll.  Guard.,  Dec.  1896,  1170;  Difficulties  in  Deep  Mines, 
M.  &  M.,  Vol.  XXII,  58. 

Laying  Out  Shaft  Bottoms,  W.  Stewart,  Coll.  Eng.,  Dec.  1896,  188;  Shaft 
Pillars,  W.  Stewart,  Coll.  Eng.,  Dec.  1896,  189;  Shaft  Bottom,  M.  &  M., 
Vol.  XX,  406,  454. 

Renewing  Old  Shafts,  Coll.  Guard.,  Vol.  LXXXI,  1991. 


CHAPTER  II. 

SINKING  IN  RUNNING  GROUND. 

Excluding  Watery  Strata. — In  he  search  for  minerals,  mining 
is  conducted  to  greater  depths  and  into  more  treacherous  ground, 
as  time  advances,  and  the  highest  type  of  engineering  skill  is 
called  into  play  as  water-bearing  strata  are  encountered.  Not 
only  is  the  shaft  to  be  sunk  through  them,  but  their  underground 
currents  must  be  excluded  from  the  mine  by  water-tight  casing, 
otherwise  an  elaborate  sy  tern  of  pumps  must  be  continually 
maintained  during  sinking  and  mining.  Not  infrequently  the 
expense  of  removing  seepage-water  while  sinking  becomes  so 
large  an  item  of  dead  work  as  to  compel  the  abandonment  of 
the  works.  A  system  of  tubbing  is,  under  those  circumstances, 
advisable. 

Two  varieties  of  cases  present  themselves — one  in  which  the 
ground  penetrated  is  quite  firm  but  porous,  and  the  other  includes 
running  soil,  marl,  quicksand,  etc.  In  the  first  case,  the  pumping 
facilities  must  be  ample,  or  the  water  kept  back  during  mining; 
in  the  second,  the  possibilities  of  excavation  are  greater  than  the 
facilities  for  its  hoisting,  and  the  laborers  are  in  danger  of  being 
overwhelmed  with  soil.  In  either  case  the  ground  traversed 
must  be  insulated  by  a  tubbing,  hermetically  sealed  above  as 
well  as  below  the  soft  measures. 

This  not  only  renders  sinking  possible,  but  it  excludes  water 
and  silt  from  the  mine,  and  permanently  dispenses  with  much 
of  the  pumping  arrangements.  It  very  seldom  comes  into  play 
in  vein-mines,  where,  with  the  vertically  of  the  lodes,  water 
cannot  be  prevented  from  percolating  into  the  mine.  It  is  in 

534 


SINKING  IN  RUNNING  GROUND.  S3S 

the  stratified  regions  that  the  use  of  the  crib  is  of  the  highest 
importance.  Beds  of  gravel,  sand,  and  clay,  or  porous  strata, 
percolating  large  quantities  of  water,  are  not  easily  traversed  or 
held  up  without  a  strong  water-tight  lining,  for  the  pressure  of 
the  moving  material  tends  to  make  the  bottom  rise,  as  well  as 
threaten  the  sides.  A  deep  shaft  in  such  a  region  may  encounter 
several  occasions  for  such  tubs,  which  under  suitable  conditions 
may  be  introduced  in  lengths  as  required,  and  only  to  the  extent 
of  the  soft  ground.  Still,  it  would  give  more  substantiality  to  the 
work  to  form  one  continuous  length  of  tubbing,  even  across  the 
good  ground.  It  is  not  uncommon  to  find  in  Germany  shafts 
with  three  sections  of  iron  tubbing,  united  by  lengths  of  brick  or 
wood  lining. 

Tubbing. — This  process  consists  in  confining  the  seepage 
area  to  the  bottom  of  the  shaft  only,  by  building  a  water-tight 
cylinder  lining  to  the  shaft,  and  carrying  it  down  with  the  sink- 
ing beyond  the  wet  stratum.  In  England,  a  bed  of  sand  called 
the  lower  red  sandstone,  which  is  almost  fluid,  has  several 
shafts  tubbed  through  it.  In  Britain,  Belgium,  and  the  North 
of  France,  several  mines  are  reached  by  tubbing  through  the 
fissured  chalks  and  marls  of  the  Cretaceous.  The  Thonmer- 
gel  of  Germany  is  frequently  tubbed  to  the  Bohn  Erz,  below, 
dry  enough  for  work.  While  sinking  the  Murton  pits  4000 
gallons  were  pumped  per  minute,  and  the  "come  in"  of  water 
for  the  Exhall  shaft  was  1650  gallons.  Still,  the  inefficiency  of 
this  plan,  sometimes  called  the  English  system,  is  recognized, 
and  several  methods  bet  er  applicable  to  loose  and  watery  beds 
have  been  applied  with  more  or  less  success.  Excepting  the 
Poetsch  method  of  freezing  the  ground  to  be  penetrated,  they 
are  modifications  of  the  diving-bell,  or  pneumatic  pile. 

Wrought-iron  tubes  in  segments  are  bolted  on  at  the  surface 
as  fast  as  the  lowering  proceeds,  until  the  secure,  impermeable 
bed  is  reached.  Here  a  smooth  base  is  prepared  for  one  or  more 
wedged  curbings,  behind  which  moss  or  concrete  is  rammed. 
The  tubbing  is  backed  with  rock  or  concrete  all  the  way  up,  and 
connected  with  the  next  upper  section.  The  holes  in  the  seg- 


53 6  MANUAL   OF  MINING. 

ments,  for  convenience  in  handling,  and  to  relieve  the  tubbing 
of  pressure  till  the  work  is  completed,  are  plugged  up.  The 
early  practice  of  bolting  the  segments  together  through  the  inside 
flanges  was  soon  abandoned,  and  now  the  flanges  are  outside, 
wedging,  pressure,  and  friction  keeping  them.  On  account  of 
the  curious  accidents  occurring  from  the  pressure  of  air  locked 
behind  the  tubes,  it  is  advisable  to  lay  a  pipe  to  the  surface  for 
the  gas  to  escape,  and,  similarly,  to  relieve  the  water.  A  shaft 
of  1 6  feet  diameter  was  sunk  at  a  monthly  average  of  104  feet 
with  four  shifts  of  8  and  10  men  each.  Several  Canadian  salt- 
mines, having  shafts  10  feet  6  inches  diameter  and  reached  at  a 
depth  of  1150  feet,  are  tubbed  through  260  feet  of  water-bearing 
strata,  in  sections  2  feet  high  and  f  inch  to  ^|  inch  thick. 
The  columns  rested  on  iron  curbs  with  firm  base.  The  joints 
are  calked. 

Masonry  Curbing. — This  is  a  much  cheaper  and  more  effi- 
cient wall  in  shafts  than  is  iron  tubing.  With  good  hoisting 
machinery,  three  masons  in  four-hour  shifts  finish  10  feet  per 
day  of  a  1 6-foot  shaft.  Towers  of  masonry,  resting  on  an  iron 
curb  with  a  cutting  edge,  were  built  on  at  the  surface;  while,  to 
facilitate  the  sinking,  digging  was  being  carried  on  below,  or,  if 
the  material  was  wet,  a  process  of  "bagging"  was  employed. 
When  abundant  in  size  and  quality,  wood  gives  great  satisfac- 
tion, being  elastic,  easily  laid  and  repaired — qualifications  not 
possessed  by  masonry.  Iron  offers  the  advantages  of  strength, 
combined  with  a  facility  of  handling,  which  recommend  it  for 
large  shafts  and  enormous  flow.  Though  it  is  not  possible  to 
presage  or  measure  the  pressure,  and  thus  determine  the  kind 
and  strength  of  tubs,  a  thickness  of  1 2-inch  wood,  y-inch  masonry, 
and  ^-inch  iron  may  be  suggested  as  common.  As  a  matter  of 
fact,  the  tubbing  should  taper  off  toward  the  top.  In  many 
cases,  however,  the  use  of  12  inch  staves  hooped  with  iron  did 
not  prove  adequate,  nor  did  the  backing  of  12  inches  more  of 
concrete  help  matters;  where  the  shafts  were  not  abandoned, 
A -inch  sheeting  met  the  emergency. 

Sinking  through  Running  Ground. — When  a  bed  is  encoun- 


SINKING  IN   RUNNING  GROUND.  537 

tered  of  a  material  so  soft  as  to  behave  like  a  fluid  and  be  pump- 
able,  the  area  which  can  be  opened  with  safety  is  very  small 
indeed.  In  this  event  some  variation  of  the  "spilling"  proc- 
esses, a  pneumatic  pile  or  some  process  of  congelation,  should  be 
employed. 

Triger's  Method  by  Pneumatic  Tube. — M.  Triger  employed, 
in  1839,  the  principle  of  the  pneumatic  pile,  in  which  an  iron  tub- 
bing was  built  down  to  an  air-lock  which  communicated  with  a 
diving-bell  at  the  bottom.  The  atmosphere  of  the  caisson  was 
maintained  at  a  pressure  of  not  over  60  Ibs.  per  square  inch, 
and  checked  the  influx  of  the  sand,  which  the  miner  shovelled 
to  a  sump,  whence  it  was  aspirated  to  the  surface.  Meanwhile 
the  tubbing  was  being  forced  down  from  the  surface  as  rapidly 
as  possible.  The  tubbing  was  divided  into  three  chambers  by 
partitions  with  hinged  doors,  the  middle  one  constituting  an  air- 
lock. The  natural  limitations  of  this  process  are  the  outside 
earth  friction  and  the  physiological  difficulties  of  men  working 
in  compressed  air. 

The  Kind  and  Chaudron  Process  of  Boring  Shafts. — In 
France  and  Germany  the  loose,  watery  marls  presented  diffi- 
culties which  the  methods  described  failed  to  overcome.  What 
with  pebbles  and  fine  rock  interfering  with  aspiration,  water 
completely  inundating  the  shafts,  and  the  difficulty  in  establish- 
ing water-tight  joints,  the  operators  were  routed.  In  1850  Herr 
Kind  devised  a  scheme  for  mechanically  sinking  shafts,  just  as 
one  does  a  bore-hole,  and  still  further  conquered  difficulties 
hitherto  insurmountable  by  a  variation  in  the  mode  of  lowering 
the  tubbing,  and  by  a  device  for  regulating  the  influx  of  water. 
When  M.  Chaudron  added  the  sliding  bottom-piece  to  form  a 
perfect  joint,  after  the  Kind  boring-tool  had  prepared  the  base, 
the  acme  of  shaft-sinking  was  reached.  Since  1862,  when  the 
first  shaft  was  sunk,  6  feet  wide,  480  feet  deep,  at  a  cost  of  $450 
per  foot,  not  a  single  fatal  accident  is  recorded  against  the  process, 
which  owes  much  of  its  success  to  the  fact  that  the  sinking  and 
lining  are  completed  before  a  man  enters  the  shaft.  Two  aban- 
doned shafts,  through  soil  feeding  11,000  gallons  per  minute, 


538 


MANUAL  OF  MINING. 


were  carried  down  267  and  216  feet,  respectively,  in  23  and  20 
months,  with  a  cost  of  $280  and  $340  per  foot.  In  the  latest 
application  569  feet  were  sunk,  16  feet  diameter,  at  an  average 
cost  of  $143  per  foot  for  both  shaft  and  lining,  which  latter  cost 
$70  per  foot.  With  a  guaranteed  success  at  so  low  a  rate,  it  is 
surprising  that  American  engineers,  usually  so  progressive,  have 
not  employed  this  method  before  acknowledging  failures;  but 
no  attempt  to  introduce  this  plan  here  is  as  yet  recorded. 

The  Trepans. — The  ordinary  procedure  comprises  first  drilling 
a  guiding  bore-pit  about  50  feet  deep  and  4.5  feet  across;  then 
widening  the  shaft  by  a  reamer  to  the  required  diameter,  and 
alternating  these  drills  with  every  50  feet  of  drilling.  These 
drills,  or  "trepans,"  are  operated  from  a  walking-beam  by  a 
surface  engine. 

The  small  trepan  (Fig.  196)  consists  of  a  blade  of  forged 
iron,  into  the  lower  side  of  which  are  keyed  a  number  of  pointed 


steel  teeth,  and  a  stem  connecting  the  blade  to 
the  suspension- rods  by  means  of  a  sliding-box. 
This  last  partially  corresponds  to  the  "jars" 
(Fig.  277)  of  oil-well  outfits,  takes  up  the  jar, 
and  is  an  essential  element  of  the  tool.  The 
trepans  are  massive — for  hard  rock,  weighing 
from  8  tons  up— and  are  raised  6  inches  or  so, 
turned  slightly  for  each  blow,  and  dropped; 
their  concussion  disintegrates  the  rock  along  a 
diameter  of  the  circle.  The  progress  in  flint  is 
3  inches  per  day;  in  chalk,  3  feet;  in  sandstone, 
i  foot,  and  in  coal  measures,  16  inches  Most  of 
the  material  is  broken  quite  fine,  though  2-inch  and  3-inch  stuff 
is  not  unusual. 

The  larger  trepan  replaces  the  smaller  one  after  50  feet  of 
drilling.  It  is  similar  (Fig.  197)  to  the  smaller  one,  but  the  blade 
is  deeper  at  the  centre  than  at  the  ends,  so  that  its  teeth  cut  a 
base  sloping  to  the  centre.  The  central,  toothless  portion  of 
the  tool  has  a  U-shaped  guide  that  fits  the  smaller  hole.  This  tool, 
often  weighing  as  much  as  16  tons,  cuts  the  shaft  to  full  width , 


FIG.  196. 
A  Trepan. 


SINKING  IN  RUNNING  GROUND. 


539 


or  it  may  be  succeeded  by  another  similar  reamer,  the  detritus 
falling  into  the  smaller  hole,  from  which  it  is  hoisted  by  the 
sludger  (Fig.  199).  In  alternate  stages  the  drilling  and  widening 
progresses,  while  the  tubbing  is  subsequently  lowered  by  a  sepa- 
rate engine.  All  these  operations  being  conducted  under  water, 
the  trepan  requires  to  be  automatically  kept  vertical.  Two 
guides,  carrying  at  the  extremities  horizontal  and  vertical  cutters, 
accomplish  this  marvellously  well.  The  record  of  the  preliminary 
shaft,  4!  feet  in  diameter,  showed  for  508  feet  an  average  progress 
of  3.3  feet  per  24  hours,  divided  up  as  follows:  51  per  cent  o£ 


FIG.  197.  Fir,.  199. 

The  Kind-Chaudron  Trepans. 


FIG.  200.. 


the  time  was  occupied  in  drilling,  19  per  cent  in  raising  and  lower- 
ing the  tools,  20  per  cent  in  dredging,  and  10  per  cent  in  repairs 
and  delays.  Widening  the  shaft  to  14  feet  and  down  460  feet 
took  ten  months;  reaming,  46  per  cent  of  the  time;  altering 
and  operating  the  tools,  etc.,  14  per  cent;  dredging,  22  per  cent;, 
delays  and  repairs,  18  per  cent. 


540 


MANUAL  OF  MINING. 


The  widening  trepan  reamer  and  sand-bucket  dredge  are 
shown  in  Figs.  197  and  198. 

There  are  no  screws  or  nuts  to  loosen;  all  the  parts  are  keyed 
to  place;  but  special  tools  are  provided  for  grappling  broken 
rods,  stems,  trepans,  teeth,  etc. 

At  the  surface  the  operations  of  boring  the  pits,  building 
and  lowering  the  tubbing,  puddling  and  sealing  the  base,  are 
conducted  with  engines  and  capstans  from  a  tall  derrick,  to 
which  extra  lengths  of  rods  may  be  attached  with  the  progress 
of  the  drilling.  Nine  men  are  employed  about  the  works,  only 
three  of  whom  are  skilled  laborers.  The  cost  of  the  installation 
of  the  machines,  tools,  etc.,  all  of  which  are  portable,  is  about 
$13,000  to  $20,000. 

The  Tubbing  and  the  False  Bottom. — This  element  of  the 
system  is  of  iron  sheeting,  built  on  in  6-foot  sections,  with  leaded 
joints,  and  suspended  by  rods.  The  flanges  are  on  the  inside  of 
the  tubbing  bb,  leaving  a  perfectly  smooth  exterior,  and  the  joints 
are  true  and  bolted  together.  Two  sections  are  lowered  daily. 
As  an  example,  a  tubbing  12  feet  7  inches 
in  internal  diameter  and  280  feet  high  is 
quoted;  it  was  i  inch  thick  at  the  top, 
if  inches  at  the  bottom,  and  weighed  400 
tons.  The  sections  were  5  feet  high,  the 
flanges  3^  inches  wide,  2  inches  thick, 
having  leaden  wedges  between,  4^  inches 
wide  and  £  inch  thick,  and  20  bolts 
i  A  inches  in  diameter. 

At  the  bottom  of  the  iron  cylinder 
are  attached  two  very  ingenious  appli- 
ances, which,  operating  automatically, 
FIG.  201. -The  False  Bottom.  have  established  the  process  as  a  success 
beyond  all  cavil;  the  first  is  a  moss-box,  a,  for  hermetically  seal- 
ing the  lower  end  of  the  tubbing  against  any  influx  of  water; 
and  the  second  is  the  introduction  of  a  false  bottom,  /,  by  which 
the  sinking  of  the  tubbing  is  cleverly  controlled.  These  are  both 
adapted  to  the  bottom  of  the  tubbing,  as  is  illustrated  in  Fig.  201. 
All  the  flanges  of  the  tubbing  turn  inward  except  the  lower  one 


SINKING  IN  RUNNING  GROUND.  541 

of  the  bottom  section,  bb,  which  is  outside,  and  may  act  as  an  an- 
nular piston  to  a  lower  section,  aa,  of  smaller  diameter,  the  upper 
flange  of  which  turns  inward,  and  the  lower  one  outward.  Be- 
tween these  flanges,  the  moss-box,  and  the  rock,  the  annular 
space  is  filled  with  moss,  which  is  not,  however,  under  com- 
pression so  long  as  the  screw-bolts  55  support  it  from  the  tubbing. 
It  operates  like  the  seed-bag  of  oil-well  diggings. 

The  false  bottom  //  is  attached  to  the  tubbing,  with  the  lower 
sections  of  which  it  forms  a  diving-bell  that  floats  the  whole 
system.  The  greater  the  head  of  water  encountered,  the  more 
complete  the  balance,  and  the  greater  is  the  relief  to  the  rods,  dd, 
supporting  the  hundreds  of  tons  of  iron.  The  safety-pipe,  g, 
with  cocks  and  plugs  operated  from  above,  is  an  equilibrium 
column  that  permits  sinking,  or  rather  regulates  its  speed. 
Opened  at  the  top,  sinking  proceeds  rapidly,  as  the  compressed 
air  and  water  find  vent;  closed,  the  whole  structure  is  upheld 
against  gravity.  When  the  plugs  are  opened  they  discharge 
into  the  tubbing,  weight  it  with  water,  and  at  the  same  time 
release  the  pressure  below.  By  proper  manipulation,  perfect 
control  is  had  over  the  lowering  of  the  casing. 

Making  a  Water-tight  Joint  at  the  Bottom. — When  the  tubbing 
has  traversed  to  the  required  depth  through  the  water-bearing 
measures,  a  seat  having  been  scraped  for  the  moss-box,  the  entire 
weight  is  allowed  to  fall  on  the  annular  piston,  b,  by  opening 
g  at  the  surface.  The  moss  is  compressed  to  a 
hard,  water-tight  mass,  the  rods  5  gliding  in 
their  bearings  (Fig.  202)  Up  to  this  time  the 
shaft  is  more  or  less  full  of  water  (the  process  is 
independent  of  the  amount  of  water  encoun- 
tered). This  may  now  be  pumped,  but  usually 
is  not  until  a  cement  backing  has  been  inserted 
and  hardened  to  insure  solidity;  after  that,  if 
the  joints  of  the  metallic  column  are  well  made, 
the  shaft  is  perfectly  tight,  and  the  mine  is  in- 
sulated from  the  subterranean  current.  The  FlG-  202« . 

The  Moss  Joint, 
introduction    of    the    cement    is    effected    by    a 

closed  spoon  holding  a  barrel  or  so,  curved  to  suit  the  space. 


542  MANUAL  OF  MINING. 

Three  sets  of  six  men  each  do  this  work,  burying  400  cu.  ft.  per 
day,  at  a  cost  of  about  forty  cents  per  square  foot  area  of  lining. 
A  solid  foundation  of  wedged  iron  curbing  is  subsequently  built 
on  a  stout  ledge,  to  take  the  weight  of  the  cribbing  after  the  other 
work  has  been  completed. 

This  method  is  generally  applicable  to  conditions  of  soft 
ground,  and  especially  in  watery  ground,  which  can  be  pierced 
without  recourse  to  ponderous  pumping  machinery.  Though 
the  pressure  of  the  water  is  not  essential  to  success,  it  materially 
facilitates  operations.  In  a  few  cases,  where  the  ground  was 
merely  wet,  not  running,  tubbing  was  cheaper  than  this  method. 
But  the  facts  should  not  be  lost  sight  of,  that  none  of  the  delays, 
perils,  and  discomfitures  of  the  ordinary  methods  are  here  experi- 
enced. Its  progress  is  greater,  and  initially  its  maintenance 
is  cheaper  than  other  schemes  for  wet  ground,  besides  never 
having  had  a  failure,  though  no  shaft  of  over  14  feet  in  diameter 
has  yet  been  sunk  by  this  method. 

An  objection  to  the  Kind  and  Chaudron  method  is  that, 
unless  an  exact  geological  section  is  at  hand,  there  are  no  means 
of  knowing  when  the  water-bearing  stratum  has  been  penetrated. 

A  short  while  ago  a  pair  of  shafts  were  sunk  in  Samlund,  East- 
ern Prussia,  for  amber,  through  147  feet  of  clay  and  sand,  by  a 
variation  of  this  method.  The  drill- tools,  weighing  1700  Ibs., 
cut  a  4-foot  6-inch  space,  though  they  had  little  to  do  except  in 
the  shale-beds.  No  moss  was  necessary,  as  the  ground  was  not 
wet.  Four-feet  lengths  of  tubbing  were  forced  down  by  jack- 
screws,  each  shift  with  27  men.  The  total  weight  of  tubbing  in 
each  shaft  was  45  tons,  and  the  total  cost  $17,500. 

Lippmann's  Drill. — This  is  a  drill  of  a  double  V-shape, 
instead  of  a  straight  trepan.  It  does  faster  work,  as  it  cuts 
equally  at  the  periphery  with  the  centre  of  the  circle.  With 
Kind's  trepan  the  blows  fell  too  far  apart  at  the  outer  edge  of 
the  shaft,  and  too  near  together  at  the  centre. 

Haase's  System. — This  employs  sheeting-piles,  in  small 
round  iron  cylinders  driven  close  together  to  form  a  cribbing 
for  the  intended  shaft.  The  tubes  were  about  15  feet  long, 


SINKING  IN  RUNNING  GROUND.  543 

^  inch  thick,  and  4  inches  in  diameter,  and  enclosed  an  area 
10X7,  which  was  then  timbered.  These  were  driven  through 
90  feet  of  quicksand  in  five  months'  time,  at  a  cost  of  $135 
per  foot.  While  standing,  they  gave  good  drainage,  and  d'd  not 
yield  when  the  excavation  of  the  shaft  began. 

The  Freezing  System. — For  loose  wet  alluvium,  Mr.  C. 
Poetsch  has  originated  the  novel  idea  of  freezing  the  mass  to  a 
solid  by  boring  a  great  number  of  holes  through  the  alluvium 
about  3  feet  apart,  and  lining  with  copper  tubes,  inside  which  are 
smaller  tubes.  A  concentrated  solution  of  the  chlorides  of  mag- 
nesium and  calcium  circulates  through  the  tubes  and  freezes 
the  ground,  after  which  the  pit  can  be  excavated  in  the  centre 
of  the  mass  in  the  ordinary  manner,  and  the  tubbing  put  in. 
This  refrigeration  is  continued  till  solid  rock  is  reached.  A 
shaft  7'Xn'  was  sunk  through  .26  feet  of  quicksand,  the  frozen 
wall  enveloping  it  being  7  feet  thick,  at  a  cost  of  $190  per 
foot. 

Conducting  the  Seepage  through  the  Walls. — Water-traps  are 
frequently  provided  in  shafts  which  are  lined  with  masonry  or 
timber.  These  consist  of  grooves  cut  around  the  entire  periph- 
ery of  the  shaft  into  a  recess  in  the  walling  of  the  shaft  and 
provided  with  pipes  and  spouts  through  which  the  water  can  be 
carried  to  some  more  convenient  point  rather  than  permit  it 
to  give  annoyance  in  the  shaft. 

If  the  waters  are  muddy  or  contain  salts  in  solution  which 
would  clog  the  spouts,  considerable  annoyance  is  given  and 
expensive  arrangements  must  be  resorted  to. 


REFERENCES. 

Artificial  Foundations  and  Methods  of  Sinking  through  Quicksand,  W. 
E.  Garforth,  Mid.  Inst,  XI,  407;  Dealing  with  Water  during  Sinking, 
Institution  of  Civil  Engineers,  Coll.  Guard.,  May  28,  1897,  994;  Sinking 
through  Quicksand,  Peter  Jeffrey,  111.  Min.  Inst.,  II,  90,  230,  240;  A  New 
Method  of  Shaft  Sinking,  G.  C.  McFarlane,  Eng.  &  Min.  Jour.,  April  7,  1900. 

Freezing  Process  Shafting,  Gobert,  Coll.  Eng.,  XVII,  171;  Freezing 
by  Compressed  Air,  M.  &  M.,  Vol.  XXII,  411;  Freezing  Water  an  Aid  to 


544  MANUAL  OF  MINING. 

Mining,  Coll.  Eng.,  XV,  86;  Sinking  of  Shaft,  B.  Barnum  Mine,  Ishpeming, 
Mich.,  R.  H.  Vondy,  The  School  of  Mines  Quarterly,  May  1882,  277;  Sink- 
ing a  Shaft  by  Compressed  Air,  Bergassessor  Luthgen,  Coll.  Guard.,  Jan.  9, 
1903;  Sinking  by  the  Freezing  Method  at  Washington,  County  Durham, 
Mark  Ford,  Ir.  &  Coal  Trds.  Rev.,  Dec.  19,  1902;  Difficulties  with  the 
Freezing  Process,  Coll.  Guard.,  Vol.  LXXXI,  201. 

Kind-Chaudron  Process,  A.  I.  M.  E.,  V,  117. 

Cleaning  Shaft  of  Water  during  Sinking,  Coll.  Guard.,  Vol.  LXXXII, 


CHAPTER  III. 

TIMBERING  ROOMS  AND  GALLERIES. 

The  Service  of  Timbering. — The  accident-tables  discussed  in 
Chapter  XIII  show  how  manifest  is  the  neglect  of  a  few  simple 
rules,  endangering  life  and  property;  and  in  no  respect  is  this  more 
painfully  impressed  than  by  the  mortality  record  of  unpropped 
rock.  Excavations,  even  in  the  "rock  of  ages,"  cannot  be  left 
open  any  great  length  of  time  without  support,  which,  if  intro- 
duced in  time,  will  prevent  disastrous  results.  Successful  super- 
intendents personally  watch  the  timbering  and  the  face-rock- 
diligently,  and  guard  against  any  springing  of  the  walls.  All 
the  effects  of  pressure  are  intensified  by  neglect,  and  the  secret 
of  success  is  to  place  timbers  before  movement  begins.  Supports- 
are  not  for  bad  roofs  only;  while  "awaiting  a  weak  spot,  the 
good  roof,  so  called,  catches  him,"  and  his  stope  or  room  is 
lost.  The  eagerness  to  quickly  win  the  face,  while  pardonable,, 
promotes  avarice,  parsimony,  want,  and  then  provokes  collapse. 

Though  the  conditions  underground  are  such  that  very  simple 
timbering  is  required  compared  with  that  on  the  surface,  the 
tendency  of  the  time  is  toward  the  employment  of  special  timber- 
men  to  make  and  place  the  supports.  In  rooms,  driving  in 
soft  ground,  and  the  like,  the  miners  must  at  once  prop  the  exca- 
vations. For  this  reason— and  the  proscribed  space— the  char- 
acter of  the  timbering  should  be  simple  and  the  sticks  light. 
Fortunately,  the  pressure  of  the  country  rock  is  inward  and  to- 
ward the  openings,  and  compression,  not  tension,  as  on  the 
surface,  is  to  be  combated  with.  This  tends  to  hold  the  sets 
together.  Tenons  and  framing  may  therefore  be  dispensed  with,. 

545 


546  MANUAL  OF  MINING. 

•except  in  loose  ground,  where  they  are  essential  for  maintaining  the 
integrity  of  the  timbers. 

The  relative  merits  of  the  different  varieties  of  wood  need 
not  be  discussed  here.  Oak  is  undoubtedly  the  most  preferable, 
but  the  mines  take  what  can  be  had  in  the  vicinity.  Above 
"timber-line"  we  are  content  with  "scrubs."  Sawn  timber  is 
better  than  hewn,  on  account  of  its  greater  resistance  to  decay; 
and  durability  is  of  prime  importance  to  strength.  Again,  green- 
wood is  heavy;  the  ordinary  lo-inch  stick,  say  6  feet  long,  is 
as  much  as  three  men  can  well  handle.  Lightness  is  an  essential 
feature  in  this  most  onerous  of  underground  work. 

The  Life  of  Timber. — This  varies  with  the  condition  of  the 
atmosphere  and  care  in  dressing.  It  is  rarely  as  great  as  that 
of  railroad-ties  (twelve  years).  In  many  mines  head-pieces 
crumble  after  two  years'  standing.  Wood  rots  faster,  and  shows 
it  less  on  the  surface  in  dry,  vitiated  air  than  in  moist  air.  Alter- 
nations of  temperature  or  moisture  are  very  destructive.  A 
cotton-fungus  mould  is  a  sure  indication  of  bad  air,  and,  being 
contagious,  requires  attention  at  once. 

The  decay  results  from  the  fermenting  of  the  albuminoids  of 
the  sap,  the  admission  of  water,  and  the  attack  of  insects,  to 
which  several  causes  contribute, — bad  air,  damp  air,  standing 
water,  and  oxidation, — causes  all  of  which  are  mitigated  by  an 
active  circulation,  and  materially  remedied  by  saturation  of  the 
pores  with  some  antiseptic.  Creosote,  Kyanizing,  or  Burnet- 
izing  will  give  greater  life  to  timber.  The  timber  is  placed  in 
a  wrought-iron  cylinder  through  end-doors,  after  closing  which  the 
air  is  exhausted  and  creosote  forced  in.  Pine  absorbs  from  10 
to  12  Ibs.  of  oil  per  cubic  foot,  and  the  hard  woods  less.  The 
pressure  during  the  operation  is  100  Ibs.  per  square  inch. 

Timber  Consumption. — While  it  cannot  be  accurately  stated 
for  the  average  mine,  it  has  been  estimated  that  the  timber  con- 
struction in  anthracite  mines  is  i  cu.  ft.  per  ton  product;  in  the 
L.  S.  copper-mines,  if;  in  Leadville  and  L.  S.  iron-mines,  3; 
.and  in  Nevada,  4^. 

Every  100  cu.  ft.  of  coal  extracted  consumes  3.4  cu.  ft.  of  tim- 


TIMBERING  ROOMS  AND  GALLERIES.  547 

ber;  every  ton  of  excavation  in  running  ground  requires  about  5 
cu.  ft.  to  support  the  balance.  In  the  Anaconda  mines,  Mont., 
80,000  cu.  ft.  of  timber  are  used  daily;  2000  of  such  mines  would 
consume  the  entire  forest-area  growth.  The  West  Vulcan  iron- 
mines,  in  L.  S.,  annually  consume  2,000,000  feet  of  lumber  and 
60,000  pieces  of  lagging,  at  a  cost  of  37  cents  per  ton  of  ore  mined. 
In  the  copper- mines  of  L.  S.  this  item  amounts  to  from  15  to  31 
cents  per  ton  of  rock  hoisted.  So  important  in  the  economy 
of  mining  and  to  the  safety  of  lives,  the  selection  and  placing 
of  timbers  should  therefore  receive  skilled  attention;  adequate 
ventilation  is  equally  urgent  for  their  preservation. 

Owing  to  its  destructibility  and  the  increasing  scarcity  of 
supply  that  is  bringing  about  the  employment  of  iron  and 
masonry  as  the  certainties  of  future  support.  Fortunately, 
the  metal-mines  above  "timber-line"  require  but  a  moderate 
supply. 

Elements  of  Timbering. — Though  the  question  as  to  whether 
the  timbering  is  to  be  done  in  a  substantial  manner  at  once,  or  to 
be  considered  as  provisional,  is  only  answered  according  to  the 
importance  of  the  gangway;  the  practice  is  to  assume  it  per- 
manent. Rooms  and  stopes  are  only  temporary,  and  treated 
as  such.  Gangways  may  be  subsequently  cribbed  or  masonried; 
but  all  pump-rooms,  machine-rooms,  and  stables  are  very  sub- 
stantially lined. 

Each  piece  should  be  placed  conformably  to  the  principles 
of  the  strength  of  materials,  and  laid  in  such  direction  as  will 
best  withstand  the  pressure  whose  direction  is  known.  The 
crushing  force  is  better  resisted  than  is  a  bending  force.  They 
should  be  placed  in  the  line  of  the  pressure  wherever  practicable, 
or  in  such  manner  as  to  act  like  or  take  the  part  of  an  arch ;  then, 
when  any  movement  takes  place r  its  effect  will  be  to  tighten  the 
timber  in  place.  Joints  should  bear  the  pressure  uniformly  and 
their  planes  be  perpendicular  to  its  direction. 

Rooms  and  levels  are  timbered  in  i-,  2-,  3-,  or  4-piece  sets. 
The  vertical  posts,  horizontal  piece,  and  top  and  bottom  sill  con- 
stitute the  usual  full  set,  and  the  inclined  stull  or  cap  represents 


548 


MANUAL   OF  MIXIXC. 


the  one-piece  or  "quarter"  set.    Usually  the  system  of  room- 
timbers  differs  from  that  in  the  levels. 

Props  and  Stulls. — Single  sticks  are  used  to  support  the 
roof  back  of  the  men.  In  long-wall  working  a  large  number  is 
used,  resting  on  the  floor  or  on  a  plate,  and  hammered  into  place 
with  a  wedge-plate  at  the  roof.  They  are  from  6  to  8  inches 
diameter,  and  stand  3  feet  apart,  in  two  or  three  rows,  beyond 
which  the  roof  caves  in  on  the  gob.  They  remain  only  a  few 
days,  are  removed  by  rows  to  let  the  roof  cave,  and  are  replaced 
nearer  the  face  (Fig.  3).  An  average  of  70  per  cent  is  recovered; 
some  of  the  balance  cannot  be  removed ;  others  \vould  endanger 
the  timbermen.  Flat  caps  on  top,  20" X  io"X2",  are  ample 


FIG.  203. — A  Stull-piece. 

for  most  bad  roofs.  Slate  requires  a  large  plate.  It  is  a  poor 
roof,  because  it  crumbles  from  the  presence  of  pyrites;  sand- 
stone or  conglomerate  makes  a  good  roof;  soapstone  is  bad; 
but  the  most  dangerous  is  fire-clay,  which  runs  when  exposed 
to  the  moist  air.  Props  come  into  play  (Fig.  210),  12  inches 
long,  3  inches  or  4  inches  in  diameter,  for  holding  up  holed  coal. 
The  props  do  not  support  the  strata  above  the  coal — this  the 
pillars  do.  They  support  only  a  portion  of  the  immediately 
overlying  seams  which  constitute  the  roof.  The  condition  of  the 
roof  and  the  method  of  mining  determines  the  number  and 
distance  apart  of  the  props.  Every  loose  block  should  be 
removed  or  propped  up. 

The  presence  of  seams  and  cleavages  traversing  one  another 
in  the  rock  materially  affects  the  selection  of  the  modes  of  tim- 


TIMBERING  ROOMS  AND  GALLERIES. 


549 


bering.  The  parallel  joints  are  not  troublesome.  Horses  in  the 
vein  usually  require  special  attention,  as  do  evidences  of  sigil- 
lariae.  The  latter  occur  like  truncated  cones,  base  down,  and  the 
circular  layer  in  the  roof  should  be  propped  as  soon  as  observed. 
It  must  be  remembered  that  the  prop,  as  the  support  for 
the  roof,  is  employed,  not  as  a  means  of  security,  but  as  a  warning 
of  excessive  pressure.  Its  elasticity  is  of  greater  importance  than 
its  strength.  By  its  bending  will  be  indicated  the  great  pressure 


FIG.  206.  FIG.  207. 

Single -stick  Timbering  in  Veins. 

to  which  it  is  subject,  and  therefore  the  necessity  for  extra  pre- 
cautions, if  the  room  is  still  to  be  kept  open,  or  for  protecting 
the  miners,  if  the  roof  is  to  be  permitted  to  fall.  For  this  reason 
props  of  cast  iron  or  of  steel  in  working  places  are  not  recom- 
mended, however  well  they  may  serve  as  a  secure  support  for 
rock  in  the  main  travelling-ways. 

States  vary  in  their  statutory  requirements  as  to  the  nearness 
of  the  props  to  the  face,  but  15    eet  is  the  farthest  al  owed  in 


55° 


MANUAL  OF  MINTSG. 


any  coal  region.      A  distance  of  5  or  6  feet  is  ample  space  for 
the  men,  but  not  for  machine  cutters. 

In  metal-mines  the  prop  is  used  as  a  stull  (Figs.  203,  205,  206), 
resting  in  a  notch  ("hitch"),  generally  on  the  foot-wall,  unless 
the  hanging-wall  is  much  softer,  and  driven  into  place  with 
a  wedge-piece,  by  mallets.  In  veins  of  small  inclination  the 
stulls  are  normal;  otherwise  they  stand  between  normal  and 
the  vertical,  because  of  the  combined  action  of  the  wall-thrust 
and  gravity.  Its  angle  toward  the  vertical  from  the  normal  is 
about  one  fourth  the  pitch  of  the  vein.  They  are  round  or 
dressed,  and  of  a  size  and  distance  apart  dependent  upon  the 
weight  of  waste  stope-rock  to  be  upheld.  It  is  better  to  increase 


FIG.  208. — Braced  Stulls. 


FlG.  209. — Vein  Timbering. 


the  number  than  the  size,  though  their  strength  is  directly  as 
the  cube  of  their  diameter.  d3  =  o.o$hw2m  is  the  formula  for. 
calculating  the  size  of  any  round  stull  held  at  two  ends,  h  is 
the  height  of  rock  along  the  vein  in  feet;  m  the  distance  between 
the  stulls  in  feet;  iv  the  width  of  the  vein  in  feet;  d  the  diameter 


FIG.  210.— Coal  Chocks. 


of  the  stull  in  inches.  To  support  60  feet  of  stull-dirt  the  timbers 
are  f  long,  12"  diameter  and  30  inches  apart.  If  reliance  is 
to  be  placed  upon  the  stull,  it  must  be  of  ample  proportions 


TIMBERING  ROOMS  AND  GALLERIES.  5$  I 

and  have  sufficient  bearing  to  receive  the  entire  pressure  uni- 
formly. 

Mill-holes. — Steep  deposits  with  good  walls  will  require  stulls 
at  distances  apart  convenient  for  the  men  to  go  to  and  from  their 
places  of  work.  If  there  is  sufficient  waste  to  furnish  convenient 
footing  for  the  men,  no  timbering  will  be  required  except  that 
used  in  lining  the  mill  hole  and  the  stulls  over  the  drift  to  sup- 
port the  waste.  Under  each  mill-hole  is  a  plat,  framed  for  the 
gate  of  the  chute,  with  loose  boards  over  the  car-track  laid  upon 
three  horizontal  spreaders  7  feet  above  the  track,  to  afford  a 
loading  chute.  In  wide  veins  the  mill-holes  are  cribbed  and 
frequently  have  two  compartments,  one  being  timbered  for  the 
storage  and  the  other  for  the  passageway. 

If  either  wall  is  soft,  a  broad  slab  or  post  is  laid  against  it  to 
take  the  thrust  (Fig.  207). 

FIG.  211.  FIG.  212. 


Underhand  Timbering. 

Braced  Stulls. — When  the  distance  between  the  walls  is  too 
great  for  a  single  convenient-sized  stull-piece,  an  arrangement 
of  shorter  sticks,  indicated  in  Fig.  208,  is  common.  A  wedge  or 
plank  is  braced  against  the  walls  and  extends  longitudinally 
with  the  drift  to  be  covered.  This  is  not  so  good  as  Fig.  209, 
wherein  one  or  two  struts  relieve  the  stull.  Figs.  211  and 
212  are  for  flooring  in  underhand  work,  and  give  support  also 
to  the  stull-dirt  overhead.  In  Fig.  209  the  saddleback,  or  strain- 
ing-beam, carries  the  load.  Not  infrequently  the  caps  may 
be  supported  by  struts  in  flat  seams,  like  Fig.  213,  or  by  a  single 
centre-prop  in  double-track  gangways  and  slopes  (Figs.  214 


552 


MANUAL  OF  MINING. 


and  215);  but,  besides  taking  up  room,  they  are  the  cause  of 
too  many  accidents.  In  fact,  much  depends  upon  the  clever- 
ness of  the  men  in  setting  the  timbers  to  the  best  advantage. 
For  example,  a  curved  stick  is  beneficially  placed  if  used  as 
shown  in  Fig.  204.  It  then  becomes  an  arch.  The  temptation 

FIG.  213.  FIG.  214.  FIG.  215. 


The  Use  of  a  Centre-prop. 

to  cut  and  notch  and  spike  should  be  restrained,  or  the  continuity 
of  the  fibres  will  be  destroyed.  The  use  of  wedges  should  be 
avoided  unless  care  be  taken  to  distribute  the  pressure  over  the 
entire  area  of  the  post  or  stull. 


FlC.  216. — Bevelled  Joints. 

Framed  Sets. — The  props  simply  support  the  roof  or  the 
stull-dirt  and  receive  longitudinal  pressure  only.  When  the  stress 
is  also  from  the  sides,  frames  are  made  in  sets  composed  of 
a  cap  or  collar  resting  on  two  posts  or  legs  studded  on  a  sill 
or  sleeper.  The  trapezoidal  form  is  stronger  than  the  rectangular, 
and  is  equally  serviceable  for  carway.  This  form  is  susceptible 
of  many  varied  modifications  of  shape,  frame,  and  joints. 

Joints. — The  joints  employed  in  gangway  sets  where  no  move- 
ment is  expected  are  the  flush  or  butt  (Figs.  218  and  219),  cut 


TIMBERING  ROOMS  AND  GALLERIES. 


553 


with  precision.  Whether  the  sticks  be  round  or  square,  the 
joints  should  be  flat.  Never  should  a  round  cap  be  made  to  rest 
in  the  hollow  of  the  post  (Fig.  217),  for  the  fit  cannot  be  made 
perfect  nor  the  splitting  of  the  post  averted.  The  cap  should 
be  shouldered  to  bear  flat  on  the  leg  (Figs.  218  and  219).  When 
the  cap  receives  vertical  pressure  only,  its  entire  width  bears  on  the 
FIG.  217.  FIG.  218.  FIG.  219. 


Flush-joint 


legs,  as  in  Figs.  218  and  221;  if  the  pressure  is  partly  from  the 
sides,  the  joint  is  dressed  to  form  as  in  Fig.  219;  for  Fig.  220  the 
lagging  and  backing  must  be  firm.  The  prop  and  collar-joint 
(Figs  222  to  225)  are  simple  and  effective;  the  bevel-joint  is  not 
uncommon  in  mining- work;  Fig.  267  is  an  elaboration  of  it,  seen 

FIG  220. 


FIG.  221. 


FIG.  222. 


Shoulder-joints. 

in  large  tunnel- work,  but  is  a  very  injudicious  concentration  of 
pressure  at  one  point,  the  avoidance  of  which  is  the  very 
design  of  framing.  The  mortise  and  tenon  joint  is  rare  in  un- 
derground work,  except  in  pump-rooms  and  the  like.  So,  also, 
there  is  little  use  for  the  scarf-joints,  unless  perhaps  in  building 
beams  for  arch  centres.  Wedges  and  head-blocks  are  essen- 
tials in  the  tightening  of  frames  and  to  lengthen  timbers.  Their 


554 


MANUAL   OF  MINING. 


removal  eases  the  work  of  reclaiming  the  sticks  whole.    The 
joints  ought  to  be  tarred  for  effective  preservation. 

Except  as  clamps,  dogs,  staples,  bands,  and  spikes,  little 
iron  is  used  in  underground  work — perhaps  i  Ib.  for  every  100 
cu.  ft.  of  lumber  placed.  Timbers  near  to  blasting  operations 
are  often  fastened,  and  bands  used  around  them  to  prevent  split- 
ting or  derangement;  otherwise  the  use  of  iron  is  not  commended, 


FIG.  225. 


FIG.  226. 


Two-piece  Sets. 

except  as  auxiliary  fasteners,  for  centres,  etc.    Iron  rusts  rapidly 
when  in  contact  with  ligneous  matter. 

The  Dimensions  of  Sets. — Sets  are  of  a  clear  height  of  5  feet 
6  inches  or  6  feet  6  inches,  and  a  width  dependent  upon  the  num- 
ber of  compartments.  A  single  way  is  about  4  feet  wide,  though 
this  leaves  little  spare  room.  The  narrowest  heading  in  which 
miners  can  work  conveniently  is  3  feet  wide  and  5  feet  high.  A 
width  of  5  feet  at  the  bottom,  and  at  the  top  of  4  feet,  is  ample 
for  all  purposes  of  a  single  way.  A  double  way  should  be  9 


TIMBERING  ROOMS  AND  GALLERIES.  555 

feet  wide  for  men  and  cars,  and  a  three-compartment  way  a 
little  wider. 

Though  the  total  cost  per  lineal  foot  may  be  somewhat 
greater  f6r  a  wider  drift  than  for  one  smaller,  the  cost  per  cubic 
yard  of  broken  rock  is  less,  and  the  difference  in  the  cost  of  tim- 
bering is  slight,  but  the  gain  in  rapidity  of  driving  markedly 
favors  large  tunnels.  Again,  as  fully  12  per  cent  of  the  mine 
fatalities  are  from  crushing  by  cars,  not  only  should  ample  room 
be  provided  for  passageway  for  the  men,  but  numerous  niches 
for  retreat. 

In  the  standard  coal-seams  a   great  height  is  not  possible, 


FIG.  227.— Reinforced  Two-piece  Set  for  Wide  Vein. 

though  not  infrequently  the  roof  is  ripped  to  secure  mule  height. 
When  the  gangway  carries  a  ventilator-box  or  gutterway,  greater 
height  is  required,  and  the  timbering  for  the  purpose  is  illustrated 
in  Figs.  228,  230,  and  231.  The  diameter  of  he  timbers  used 
is  as  small  as  will  serve  the  purpose,  though  it  is  often  1.4 
inches.  Large  timbers  are  often  needed,  but  not  placed,  because 
inconvenient  to  handle.  Instead  of  the  building  of  thicker  pieces, 
the  sets  are  placed  nearer  than  the  average — 3  feet  6  inches; 
skin  to  skin  is  not  unusual  in  shattered  ground.  The  height  of 
slopes  depends  upon  the  mode  of  haulage  The  use  of  carriage 
requires  great  height;  with  a  dip  of  less  than  40°,  the  height  is' 
about  7  feet.  For  skips,  the  normal  height  of  6  feet  will  do 


556 


MANUAL   OF  MINING. 


Gangway  Sets. — Four  pieces  usually  constitute  the  frame, 
but  the  sill  may  be  dispensed  with  where  the  floor  is  good.  In 
slopes  sills  are  essential,  wedged  into  place  to  secure  the  track- 
way. In  laying  the  sills  the  trench  is  dug  lower  in  the  centre 
than  at  the  ends,  so  they  will  not  break.  Upon  their  ends  the 


FIG.  230. 


FIG.  231.  FIG.  232. 

Three-piece  and  Four-piece  Sets. 


posts  rest  to  support  a  cap,  which  is  wedged  into  bearing  by 
blocks  and  packing. 

Where  a  level  road  is  driven  through  firm  material,  only  the 
roof  of  which  needs  support,  the  post  and  collar  form  of  tim- 
bering (Fig.  236  or  223)  will  suffice.  In  pitching  seams  a  variety 
is  employed,  iis  in  Fig.  232,  when  the  vein  and  country  rock  are 
sound  and  the  hanging-wall  soft.  With  a  poor  roof  and  firm 
vein-  matter  one  leg  is  floored  (Fig.  227),  and  the  other  rests  on 
the  vein.  This  is  also  seen  in  coal  regions.  Fig.  233  is  a  strong 
form  for  pressure  from  sides  and  top.  For  such  conditions  the 
arched  form  of  short  timbers  may  be  utilized  as  in  Fig.  204. 

Sets  are  laid  close  together,  or  the  distance  between  them  is 


TIMBERING  ROOMS  AND  GALLERIES. 


557 


Timbering    Loose    Ground. — Should  the  vein-matter  be  too 
loose  to  stand  up,  the  gangway  is  lagged  against  a  long  brace 
FIG.  233.  FIG.  234. 


Wide-vein  Timbering. 

(Fig.  227).    In  wide,  soft  veins  a  similar  idea  is  employed,  as 
in  Fig.  234. 

Gob-roads  are  timbered  only  at  the  roof,  by  the  lagging  of 
caps,  resting  on  dry  pack-walls.     Occasional  chocks  will  give 


FIG.  235. — Timbering  a  Gob-road. 

greater  consistency  to  the  whole  (Fig.  235).  But  as  the  sub- 
sidence of  the  roof  cannot  be  prevented  the  road  cannot  be  kept 
open  without  repairs  until  the  gob  is  pressed  solid.  Meanwhile 
the  maintenance  is  a  serious  item.  The  plan  of  keeping  a  road 
open  along  the  face  of  coal  through  waste  is  never  to  be  com- 
mended. The  mode  of  timbering  alluvial  gold-mines,  called 
"blocking,"  is  said  by  the  inspector  of  mines  at  Sandhurst  to  be 
perfectly  safe.  The  prop  and  collar  system  is  used  while  pass- 
ing the  chain  pillar  on  either  side  of  the  gangway;  beyond,  a 
modification,  in  which  each  cap  is,  in  turn,  made  to  rest  upon 
the  caps  of  the  preceding  sets,  already  built,  and  upon  the  collar 


558 


MANUAL   OF  MINING. 


of  the  next.    The  timbers  are  8  inches  in  diameter.     In  very 
soft  ground  each  cap  has  two  posts  for  its  support. 

Timbering  Roads  in  Large  Deposits. — This  is  difficult.    The 
gangway  takes  only  a  small  portion  of  the  width  of  the  vein; 
the  rest,  if  firm,  is  left  to  stand,  or  if  loose,  packed  with  waste. 
FIG.  236.  FIG.  237. 


Special  Methods  of  Frames. 

In  the  Great  Devon  Consols  mine  the  compartments  are  used 
one  for  "attle"  (waste),  and  the  other  for  travelling  (Fig.  237). 
The  vein  is  22  feet  wide;  the  stulls  are  20  inches  and  18  inches. 
In  a  24-foot  vein  a  sole-piece  of  24-inch  timber  and  struts  of 
18  inches  with  a  longitudinal  piece  at  the  apex,  made  a  very 
FIG.  238.  FIG.  239. 


Reinforced  Timber  Sets. 

strong  frame  (Fig.  234).  With  3-inch  plank  lagging,  from  10  to  50 
fathoms  of  attle  are  carried.  The  hitches  are  cut  18  inches 
deep.  In  Southern  France,  with  great  pressure  from  the  roof 
and  for  supporting  heavy  waste,  Figs.  238  and  239  are  much 
seen.  In  many  cases  an  arch  of  vein-matter  10  feet  thick  remains 
untouched  from  the  stope  below.  With  a  system  of  filling  this 
arch  is  subsequently  recovered.  In  fact,  without  filling,  no  large 
deep  mine  can  be  held  by  timbers. 


TIMBERING  ROOMS  AND  GALLERIES. 


559 


FIG.  240. — Square-set  Timbering. 


56° 


MANUAL   OF  MINING. 


If  the  vein-matter  is  very  soft,  the  character  of  the  framing 
must  be  entirely  altered.  In  the  Austrian  salt-mines  the  prob- 
lem is  very  difficult,  because  the  material  assumes  the  nature  of 
a  fluid.  In  rock  that  decomposes  upon  exposure  to  the  air  the 
timbering  involves  some  elaborate  form  of  framing.  The  gang- 
ways in  a  vein  or  bed  may  then  be  a  component  part  of  the  square- 
set  system  (Fig.  240).  Where  ground  has  a  tendency  to  swell, 
the  only  way  to  save  the  timbers  is  to  ease  up  the  ground  behind 


FlG.  241. — Heavy  Timbering  for  Galleries. 

the  timbers  from  time  to  time  until  the  ground  settles  to  its  nat- 
ural state.  The  swelling  can  neither  be  prevented  nor  resisted.. 
For  shifting  ground,  the  style  of  Sutro-tunnel  timbering  (Fig. 
241)  is  preferable  to  that  illustrated  in  Fig.  268,  which  has  proven 
insecure. 

In  relieving  timbering  and  tunnel  linings  of  the  result  of  con- 
tinual movement  an  admirable  arrangement  consists  in  driving; 
a  lateral  tunnel  or  two  with  heavy  timbering,  but  having  an  open 


TIMBERING  ROOMS  AND   GALLERIES.  561 

face  to  serve  as  a  safety-valve  through  which  the  excessive 
pressure  will  be  expended. 

Few  cross-cut  tunnels  require  timbering  further  than  50 
feet  or  so  from  the  mouth,  for  the  ground  stands  well.  In  pitch- 
ing rock  and  porphyry  the  roof  should  be  heavily  braced. 

Lagging. — If  the  spaces  between  the  sets  cannot  be  left  open, 
the  sides  and  top  are  lagged  with  plank  or  "slabs"  from  the 


FIG.  242.— Iron  Lining  for  Double-track  Gangway. 

saw-mills.  These  are  driven  close — with  the  flat  side  out,  though 
they  are  better  reversed.  Lagging  shou  d  never  be  very  strong; 
the  slabs  are  always  weaker  than  the  members  of  the  frames, 
and  serve  merely  to  prevent  the  fine  rock  from  sifting  into  the 
gallery  and  leaving  an  open  space  that  gives  opportunity  for 
movement,  which  if  once  begun  can  never  be  resisted.  A 
very  slight  movement  produces  sufficient  pressure  to  break  the 


5$ 2  MANUAL  OF  MINING. 

lagging,  and  thus  relieve  the  costlier  work.  These  are  readily 
replaced  on  occasion.  In  many  cases  round  poles  are  used,  3 
or  4  inches  in  diameter,  overlapping  two  sets.  If  the  roof 
does  not  run,  a  few  logs  suffice  (Fig.  227).  In  soft  ground  or 
under  poor  roof  the  men  are  protected  and  advance  made  by 
"poling"  ahead  of  the  face  (Fig.  271).  The  open  space  between 
the  lagging  and  the  rock  is  packed  with  waste  or  wedged  per- 
fectly. Brush  piled  back  of  the  lagging  holds  up  the  "smalls" 
well. 

When  the  capping  over  the  firm  ore  of  a  flat-bed  is  not  good, 
timbering  may  be  saved  by  not  stripping  the  entire  height  of 
the  vein,  but  leaving  a  layer  of  it  for  roof  to  be  removed  later. 

The  Use  of  Iron  Underground. — Of  the  relative  merits  of 
iron,  wood,  and  mason  y  much  is  heard.  Suffice  it  to  say,  that 
in  Europe,  wrhere  the  utmost  care  is  taken  to  preserve  the  timber 
— by  replanting  for  each  tree  cut  down — iron  and  masonry 
are  put  to  extensive  service.  As  to  props  of  metal  or  of  wood,  the 
number  would  be  the  same — more  for  a  brittle,  less  for  a  flexible 
roof;  and  whatever  the  condition  of  the  roof,  the  size  of  the 
metal  props  would  be  nearly  the  same  Iron  props  are  of  the 
-f  or  O  cross  section,  set  on  the  thill  or  upon  a  foot-block,  to 
be  drawn  by  lever  or  bar  and  chain  Jack-screws  have  been 
used,  but  their  expense  is  too  great  for  their  general  use.  Cast- 
iron  props,  auxiliary  to  pack-walls,  9  feet  apart,  5  feet  long, 
4  inches  outside  diameter,  weighing  150  Ibs.,  have  been  employed 
in  collieries;  their  use  is  emphatically  stated  to  be  cheaper  than 
wood. 

Levels  are  not  infrequently  lined  with  iron  tubbing  similar 
to  that  adopted  for  shafts 

The  increased  price  of  timber  for  underground  work  has  in- 
duced the  use  of  steel.  The  common  structural  sizes  of  steel 
are  used  in  the  galleries.  The  sets  are  usually  of  the  three-piece 
lype  and  the  structure  is  of  the  strap  type  of  connection.  I  beams 
or  channels  are  used  for  the  caps  and  posts,  the  latter  being 
bevelled  for  the  batir  desired.  The  connections  are  made  by 
angle  straps,  bolted  or  riveted  into  place.  The  foot  of  the 


TIMBERING  ROOMS  AND  GALLERIES. 


563 


posts  is  usually  set  into  a  cast  shoe  upon  a  cast  base.  In  some 
mines  the  pieces  are  connected  by  pins  instead  of  angle  straps, 
the  latter  being  cheaper  but  more  difficult  in  connection.  In 
choosing  the  size  of  the  members  the  heaviest  shapes  are  usually 
taken  in  order  to  obtain  the  thickest  possible  web  and  thus  a 
small  size  of  the  pins.  When  channels  are  used  in  pairs  with 
gas-pipe  separators,  they  have  a  special  socket  for  the  foot. 

Masonry  Lining  for  Roads. — The  use  of  masonry  for  drifts 
and  tunnels  is  very  common.  Where  timber  is  rapidly  destroyed, 
or  where  the  pressure  is  too  great  for  rectangular  frames  to  be 
•economically  employed,  the  arch  is  pressed  into  service.  On  the 
other  hand,  good  stone  must  be  plenty  and  cheap.  It  is  laid 
dry  or  cemented,  with  the  walls  straight  or  curved  in  the  top 
arch.  The  principle  usually  followed  is  to  turn  the  arch  out- 
FIG.  243.  FIG.  244.  FIG.  245. 


The  Segmental  Arch  as  a  Stull. 

ward  against  the  direction  of  pressure.  The  top  and  the  floor 
may  be  arched  and  the  sides  in  four  separate  curves,  or  they  are 
combined  into  one  geometrical  figure,  either  the  egg  shape,  ellipse, 
or  circle.  In  some  mines  the  masonry  is  built  exclusively  in  all 
the  openings,  dispensing  entirely  with  the  use  of  timber  In  others 
only  the  main  underground  arteries  are  walled.  Generally  the 
masonry  is  built  after  preliminary  framing  has  served  for  the 
exploration  work.  The  walls  may  be  continuous  along  the 
entire  length  of  the  gallery  or  may  be  in  sections  where  the  wall- 
rock  is  not  strong  enough  to  dispense  with  auxiliary  support. 
The  sides  of  air-crossings  require  walling,  as  also  the  spaces 
where  ventilating-doors  are  fixed  between  their  posts  and  the  sides 
of  the  roads.  Whatever  the  form  of  the  masonry  or  lining,  it  is 


504 


MANUAL   OF  MINING. 


indispensable  that  the  spaces  back  of  them  shall  be  filled  with 
waste  rock  and  contain  no  decomposable  material. 

Where  the  roof  is  firm  and  the  sides  only  are  weak,  straight 
walls  are  built  of  stone  or  brick  irom  a  foundation  a  few  inches 


FIG.  248. 
Masonry  Arches  on  One  Soft  Wall. 

below  the  floor  line  to  the  roof  on  either  side  (Fig.  252).  If 
there  is  a  slight  pressure  from  the  sides  the  walls  may  be  inclined 
or  receive  a  batir  to  give  a  secure  base.  Upon  these  walls  timber 
caps  or  struts  covered  with  lagging  are  frequently  employed 
to  support  the  roof  instead  of  leaving  the  roof  undisturbed. 


TIMBERING  ROOMS  AND  GALLERIES. 


565 


The  arch  is  preferred  for  all  permanent  ways.  It  may  be 
segmental  or  semicircular  for  the  support  of  waste  in  stopes, 
according  to  the  weight  to  be  upheld.  They  are  seldom  more 
than  two  bricks  thick,  and  when  properly  laid  should  be  able  to 
withstand  all  the  pressure  which  emergencies  require. 

In  Figs.  246  to  248  are  illustrated  arches  when  the  vein-matter 
and  the  hanging-wall  are  too  weak  to  be  self-supporting.  Firm 


FIG.  249. — Arch  Centre. 


FIGS.  250  and  251. — Arch  Failures. 


seats  for  the  arch  must  be  provided  or  imposts  are  built,  as  in 
Fig.  247. 

Arch  Centres. — Whatever  the  procedure,  arches  are  built  on 
centres  and  by  template,  for  invert  and  walls.  The  centers 
should  be  made  of  light,  small,  easily  framed  sticks,  that  are 
not  so  close  as  to  interfere  with  work,  yet  strong  enough  to  sup- 


FIG.  252.— Walling  Gob-roads. 

port  the  thrust  that  may  fall  on  them  when  the  tunnel-timbering 
is  removed.  Its  shape  may  be  whatever  is  the  most  convenient 
for  the  traffic.  The  elliptical  linear  arch  is,  however,  the  form 
most  commonly  adopted,  the  side  and  roof  comprising  the  upper 
part  of  the  ellipse,  which  is  closed  below  by  a  segmental  invert 
arch,  with  the  springing  lines  on  horizontal  faces.  In  stratified 
rocks,  the  strongest  form  for  the  roof  is  that  of  a  pointed  arch. 


566  MANUAL   OF  MINING. 

Sometimes  in  solid  rock  the  horseshoe  form  is  used  for  the  top 
and  sides,  the  floor  being  level. 

The  two  accompanying  figures,  250  and  251,  may  be  in- 
teresting as  suggesting  the  places  of  weakness  with  the  given  con- 
ditions of  pressure.  If  the  excessive  pressure  is  from  the  top 
and  any  opportunity  for  bulging  is  given,  the  collapse  will  take 
place  as  shown  in  Fig.  251.  If  the  side  pressure  is  very  great 
and  the  roof  resistance  small,  the  break  occurs  at  the  keystone. 

Tubular  Walling  of  Galleries.  —  In  German  mines  will  be 
seen  examples  of  tubular  walling  of  the  elliptical  (Figs.  253  and 
254)  or  inverted  oval  form,  the  choice  between  them  for  greatest 
strength  being  still  an  unsettled  matter.  The  latter  gives  greater 
width  at  the  bottom  and  smaller  area  for  pres- 
sure at  the  top.  Masonry  cannot  be  built  unless 
the  ground  is  previously  timbered,  or  firm 
enough  to  stand  while  the  mortar  is  drying. 
In  soft  ground  the  level  is  driven  by  spilling, 
which  can  only  be  replaced  by  the  masonry 

FIG.  253.         retreating  toward  the  shaft.      When  the  tempo- 
Masonry  Walling. 

rary  service  of  the  timber  has  been  accomplished, 

and  masonry  is  to  be  substituted,  the  uprights  are  cut  at  the  foot, 
the  sills  and  spilling  laths  (Fig.  273)  removed,  the  bottom  arch 
is  made  first,  side  walls  next  replace  the  posts,  the  cap  being  tem- 
porarily supported  on  props,  the  centre  set  up,  and  the  top  laid. 
Figs.  255  and  256  represent  a  masonry  walling  employed  where 
the  floor  is  sound. 

Dams. — Dams  for  keeping  back  water  are  either  straight- 
backed  (Fig.  257)  or  arched. 

The  use  of  iron  is  advocated  and  receiving  ready  acceptance 
in  mining-  as  well  as  tunnel-work.  The  life  of  timber  is  short, 
its  resistance  low,  and  the  component  parts  of  the  frame  must 
be  rigidly  connected.  The  ordinary  constructive  forms  of  iron 
are  applied  in  the  ordinary  way  for  columns,  caps,  or  arches. 

Timbering  Soft  Ground. — In  soft  ground,  which  is  liable  to 
run,  some  form  of  stout  framing  is  indispensable.  Indeed,  ore- 
bodies  of  large,  irregular  dimensions  arc  peculiarly  adapted  to 


TIMBERING  ROOMS  AND  GALLERIES. 


567 


some  form  of  cribbing  or  square  sets.  The  latter  is  a  natural 
extension  of  the  system  of  running  contiguous  parallel  drifts 
in  the  ore  in  two  or  more  stories.  In  the  creviced  matter  of  the 
Comstock  mines,  in  the  very  poor  ground  of  the  Lake  Superior 
region,  in  iron-mines  where  pillars  cannot  be  trusted,  in  the 
FIG.  254.  FIG.  255.  FIG.  256. 


Stone  Walling  for  Gangways. 

rotten  lead-ores  of  the  Leadville  beds,  and  in  the  deposits  at 
Butte,  Montana,  an  extensive  system  of  framing  is  in  use  and 
has  found  ready  acceptance  in  various  sections  of  the  world.  As 
a  distinctive  method  of  mining,  it  was  referred  to  in  page  47, 
and  though  many  properties  are  substituting  a  filling  method, 
it  still  retains  a  hold  on  the  mining  fraternity  that  qualifies  it 


FIG.  257. — Dams. 

for  a  place  here.     Still,  several  causes  hasten  the  decline  of  its 
popularity. 

The  plan  is  devised  for  soft  ground,  and  is  equally  well  adapted 
to  large  openings  in  rock  with  a  small  tendency  to  cave,  slide,  or 
swell.  The  system  is  a  simple  and  natural  extension  and  refine- 
ment of  the  practice  of  running  contiguous  parallel  drifts  in  the 
ore  and  in  two  or  more  stories  when  the  thickness  of  the  ore 
required  it. 


5 68  MANUAL   OF  MINING. 

The  Square  Set. — The  square  set  in  stoping  requires  vast 
quantities  of  timber,  and  its  framing  is  also  expensive.  Never- 
theless, its  adaptability  to  all  forms  of  excavation  commends  it. 
The  timbering  of  shaft,  slope,  and  slopes  may  all  be  rigidly  con- 
nected in  one  system  with  facility.  Though  there  are  objections 
to  it,  nevertheless  it  frequently  proves  in  the.  end  to  be  cheaper 
and  more  economical.  The  open  cells  of  timber  may  be  inserted 
in  the  event  of  an  unexpected  thrust. 

The  set  is  made  up  of  posts,  cap,  girt,  and  frequently  a  brace, 
which  may  be  the  middle  number  of  a  set  of  three  pieces  form- 
ing the  letter  N  inside  of  the  main  set.  The  cap  usually  is 
placed  across  the  deposit,  while  the  girt,  also  resting  on  the  post, 
extends  longitudinally  with  the  roof  of  the  ore-body.  They  are 
usually  framed  for  7  feet  of  clear  height  to  allow  of  reinforce- 
ment sets  being  placed  later,  and  still  to  leave  ample  passageway. 
The  posts  are  5  feet  apart  in  the  clear. 

The  usual  plan  consists  in  driving  the  heading  along  the 
level  and  near  the  centre  of  the  deposit,  from  which  is  extended 
the  timbering.  Two  sill-pieces  and  girts  are  placed,  then  the 
posts;  after  which  a  cap  across  the  drift  and  girts  connects  this 
set  with  the  one  last  placed  in  position.  The  frame  is  block?d, 
plumbed,  and  wedged  against  the  back  of  the  drift.  As  fast 
as  excavation  proceeds  on  either  side  of  the  heading,  one- post 
sills  are  laid  on  the  floor  and  penned  to  the  two-post  sills,  a  stull- 
girt,  then  a  post  fitted  on  that,  the  cap  and  girt  holding  its  top 
in  position.  As  soon  as  wedged  against  the  top  and  the  side,  two- 
inch  plank  roofing  is  laid  from  cap  to  cap  on  top  as  a  floor  for 
the  upper  set. 

This  continues  to  the  foot-wall,  along  which  the  sets  are  car- 
ried by  means  of  a  cap-sill,  which  combines  the  functions  of  the 
two  pieces  as  shown  in  the  illustration.  From  this  a  new  line  of 
sets  is  carried  up  indefinitely. 

Thus  each  one  of  the  full  square  sets  encloses  a  rectangular 
volume  or  cell.  Each  post  supports  two  cross-sills  and  two  caps 
and  rests  upon  four  sills  (Fig.  258).  Note  also  in  Chapter  III, 
Part  I,  the  method  of  square-set  timbering. 


TIMBERING  ROOMS  AND  GALLERIES. 


569 


The  frames  are  built  up  as  fast  as  the  work  is  opened,  unless 
it  happen  that  the  ground  will  stand  a  while,  until  the  timbermen 
can  attend  to  the  chamber.  In  ore  that  will  not  remain  in  place 
long  enough  to  advance  a  set,  an  intermediate  false  set  of  cap 
and  uprights  supports  lagging  overhead  until  the  men  reach  the 
full  length  of  a  cut.  If  the  ground  will  not  allow  of  this  advance, 
it  runs,  and  only  caving  or  filling  is  admissible. 


FIG.  258. — Sc 


The  method  of  procedure  varies  in  different  regions,  or  per- 
haps with  the  nature  of  the  ore.  The  lower  floor  is  worked  out 
first,  the  bents  being  added  on  at  the  right,  left,  and  ahead; 
after  which  the  next  tier  is  set  in  the  same  manner  directly  over 
the  first.  A  species  of  overhand  stoping  is  sometimes  employed 
whereby  the  floor  tier  is  progressed  only  a  set  or  two  before  the 
next  upper  is  advanced,  to  be  followed  later  by  another  set  above, 
and  so  on  up.  The  order  is  a  matter  of  indifference  provided  a 


57°  MANUAL  OF  MINING. 

perfect  alignment  is  secured.  Where  this  method  is  used  in 
steep  veins,  very  careful  surveying  is  necessary  to  carry  up  the 
tiers  of  the  lower  level  to  a  line  with  those  in  the  upper  stope. 

When  a  sill-floor  is  to  be  laid  upon  a  large  body  of  ore  to  be 
mined  from  the  next  lower  level,  the  sills  should  cover  several 
sets  to  render  less  dangerous  the  work  of  connecting  the  lower  nest 
of  timber  with  the  upper.  The  plates  against  the  walls  should 
also  cover  several  sets,  if  the  walls  are  bad  or  crumbly. 

Sticks  10  inches  in  diameter  or  square  are  used  for  7 -foot 
sets,  and  some  as  large  as  14  inches  are  not  uncommon  with 
9-foot  posts.  They  are  dressed  to  a  faultless  fit,  being  cut  to 
template  by  some  saw  like  the  Hendey  (Fig.  260). 

When  the  permanent  roof  is  reached,  lagging  is  laid,  wedged, 
and  packed. 

Reinforcing  Square  Sets. — Where  the  posts  cannot  be  placed 
in  the  direction  of  the  greatest  pressure,  reinforcing  sets  are  used 
in  heavy  ground,  with  the  planes  of  the  inclined  piece  in  the 
direction  of  the  greatest  pressure.  The  sets  then  assume  the  form 
of  X  or  N. 

Objections  to  the  System. — All  the  sets  are  dependent  upon 
one  another,  and  therefore  an  absolutely  perfect  fit  is  essential 
to  transmit  the  pressures  equally  and  to  maintain  the  framing 
intact.  If  the  joints  are  not  true,  each  one  will  have  a  slight 
play  and  an  opportunity  for  movement  will  be  afforded.  The 
caps  may  slip  off  the  posts.  The  sets  should  be  "plumbed" 
occasionally  to  watch  for  incipient  displacement.  Sometimes,  to 
prevent  this  form  of  failure,  solid  cribs  of  timber,  two  sets  wide> 
are  built  across  the  ore-body  from  one  wrall  to  the  other. 

The  sets  may  crush  when  the  safety-valves  of  lagging  have 
been  omitted.  These  consist  of  frames  or  lagging,  slightly 
weaker  than  the  timbering,  arranged  to  be  easily  replaced.  Fel- 
site  or  trachyte  tends  to  swell  and  crush  timbers,  and  in  such 
ground  should  be  closely  watched.  At  the  first  sign  of  fracture 
of  any  member  of  the  system  it  is  advisable  to  at  once  withdraw 
the  men  from  that  locality;  for  ruin  spreads  very  rapidly  after 
the  destruction  of  a  single  member  of  a  framing  so  loosely  con- 


TIMBERING  ROOMS  AND  GALLERIES. 


57* 


572 


MANUAL   OF  MINING. 


nected  as  is  the  square  set.  A  chamber  90  feet  high,  with  13 
tiers  of  timbers,  was  in  complete  ruin  within  thirty  minutes  after 
the  first  sign  of  break  of  a  sill. 

Cribs. — Rooms  and  abandoned  large  stopes  are  supported 
by  massive  columns  of  heavy  timbers  carried  to  the  roof  and 
filled  with  waste.  These  "cribs"  are  also  built  as  an  abut- 
ment in  a  chamber  having  the  wall  as  a  back.  A  good  founda- 
tion is  prepared,  of  the  proper  size  and  shape;  two  logs  are 
laid  parallel,  and  upon  them,  in  the  notches  at  the  ends  and  at 


FIG.  260. — A  Crib. 

the  middle,  three  cross-sills  are  laid;  upon  them  again  rest  a 
pair  of  sills  slightly  inside  and  above  the  others;  upon  these, 
in  turn,  another  layer,  etc.  (Fig.  260).  Inside  of  this  space,  as 
fast  as  enclosed,  waste  is  piled.  The  logs  are  10  feet  and  14  feet 
long,  and  12  inches  to  20  inches  in  diameter  at  the  butt.  Crib- 
bing may  also  be  built  with  the  removal  of  the  ore  in  overhand 
stoping,  the  waste  being  utilized  to  fill  the  crib.  As  the  height  of 


TIMBERING  ROOMS  AND   GALLERIES.  S73 

these  cribs  increases,  the  area  of  their  base  is  proportionately 
increased. 

These  pens,  or  cribs,  are  also  used  for  construction  of  dams  or 
abutments  at  mine-dumps,  the  front  face  consisting  of  long 
face-sills  parallel  with  face  of  the  hill,  with  cross-sills  projecting 
from  the  hillside  and  resting  on  the  face- sills.  They  may,  or, 
may  not,  be  notched.  The  structure  continues  as  high  as  desired 
the  framing  being  carried  up  as  fast  as  the  waste  rock  accumu- 
lates upon  those  already  laid  in  position. 

Mill-holes  are  built  of  cribs  30  inches  square  in  one  or  two 
compartments. 

It  is  questionable  if  there  is  any  choice  between  the  square 
set  and  cribbing  in  large  rooms.  The  former  is  well  suited  to 
turning  off  into  small  cells  at  the  ends  of  the  room,  but  it  is  danger- 
ous if  side  pressure  exists.  Cribbing  will  never  do  in  very  soft 
ore  under  brittle  roof.  This  arrangement  constitutes  a  very 
strong  "made"  dump,  where  the  rock  is  not  permitted  to  roll 
away  freely,  as,  for  instance,  on  account  of  contiguous  buildings 
down  the  hill. 

Protecting  Underground  Chambers. — Underground  chambers 
intended  for  steam-pumps,  hoisters,  stable,  etc.,  are  built  in  any 
suitable  shape  that  provides  sufficient  room,  and,  being  large, 
require  great  skill  to  utilize  framing  or  walling  materials  to  the 
"best  advantage.  Undoubtedly  an  arched  roof  will  give  the 
greatest  resistance  and  strength,  and  masonry  is  therefore  sug- 
gested. Besides,  the  hot,  damp  atmosphere  of  steam  would 
rapidly  destroy  timber.  Still,  the  latter  is  more  convenient  than 
masonry,  in  heavy  sets  of  three-piece  or  five-piece  arches,  and 
spliced.  They  are  laid  close  together,  lagged  over,  and  packed 
to  prevent  wedging  apart.  The  ventilation  of  these  rooms 
should  receive  special  attention.  For  air- compressors,  engines, 
and  coal-cutters  an  enlarged  level  will  do,  with  some  stouter  caps, 
railroad  rails,  and  I  beams  on  the  wall-posts. 

Timbering  Landings. — The  timbering  plats  and  landings 
must  vary  with  the  character  of  the  intersections.  The  frames 
must  support  one  another  as  well  as  the  country  rock,  and  should 


574  MANUAL   OF  MIX  IXC. 

afford  firm  fastenings  for  the  plats  and  doors  used  for  the  landing 
of  the  cars  and  buckets.  The  level  or  gallery  should  be  widened 
near  the  shaft,  to  give  room  for  sidings,  storage  closet  for  powder 
and  steel. 

The  masonry  lining  of  the  shaft  is  supported  by  an  arch  to, 
or  by  lintel  on,  two  posts  or  walls  at  the  sides  of  the  gallery  land- 
ing. The  former  gives  a  high  opening  for  landing  the  buckets. 

Mill-holes  carried  up  with  the  waste  are  solid  cribs  of  30  inches 
square  for  a  manhole,  and  may  also  have  a  compartment  for 
sliding  ore  (Fig.  87).  Either  cordwood  or  sawed  blocks  are 
used  for  the  lining.  The  latter  plan  gives  them  greater  dura- 
bility, for  the  abrasion  is  along  their  cut  faces. 

The  timbering  of  slopes  is  similar  to  that  of  galleries,  except 
that  greater  care  is  taken  in  cutting. 

Timber-cutting  Tools. — The  timberman's  tools  are  few.  They 
need  be  only  hatchet,  hammer,  and  wedge,  with  a  bar  and  chain 
for  casual  work.  The  timber  should  be  delivered  below  ready 
for  insertion.  A  sawmill  at  the  surface  is  now  as  much  a  com- 
ponent of  the  surface  improvements  as  is  the  boiler. 

Mine-timbsr  Framing-machine. — In  Fig.  29  is  illustrated  a 
machine  which  is  designed  to  saw  tenons  upon  timbers  of  stand- 
ard size.  It  consists  of  a  number  of  adjustable  saws  which 
may  be  arranged  at  any  distance  apart  to  produce  the  desired 
dimensions  of  the  head  on  the  timbers.  By  a  machine  set  to  a 
given  pattern,  all  timbers  are  finished  to  an  exact  shape,  which 
when  placed  in  position  ensures  an  alignment  of  the  framing, 
the  joints  being  accurate  and  the  contact  perfect.  Security  is 
thus  attained  and  the  life  of  the  frame  is  prolonged. 

Though  the  work  of  timbermen  cannot  accurately  be  stated, 
the  following  is  given  as  the  experience  of  an  "old  miner"  for 
one  day's  labor: 

Making  about  TOO  feet  of  wooden  ladders; 
Placing  65  sq.  ft.  of  close  shaft-lining  down  to  a  depth  of  70  ft.; 

40  "    "    "     "          r    "  "      "  "       "     "200  " 

Cutting  and  dressing  mine  frames,  50  running  feet; 
Framing  mine-sets,  40       ' ' 


TIMBERING  ROOMS  AND   GALLERIES.  575 

Dressing  sills,  30  running  feet; 
Making  60  poling  planks; 
"    100       "      wedges; 
Setting  up  20  pairs  of  props  and  head-blocks. 

REFERENCES. 

Timbering,  M.  &  M.,  Vol.  XXIII,  117;  Theory  of  Timbering,  M.  &  M., 
Vol.  XX,  495;  XXII,  566;  Min.  Inst.,  X,  726;  Dams,  M.  &  M.,  Vol.  XX, 
445;  Structural  Timber,  Comparative  Tests  of  Bracing,  Prof.  E.  Kidwell, 
L.  Sup.  Min.  Inst.,  IV,  1896,  34;  Timbering  of  Silver  and  Gold  Mines,  J.  T. 
Freeland,  Coll.  Eng.,  1895,  100;  Timbering:  Economical  Use  of  Timber  in 
Mines,  H.  W.  Halbaum,  Coll.  Guard.,  Aug.  1896,  407;  Mine  Timbering,  J. 
Clark  Jefferson,  Chest.  Inst.,  VII,  270;  Mine  Timbering,  C.  W.  Swallow, 
Sci.  Am.  Sup.,  Sept.  8,  1894;  Timber  in  Mining  Regions  about  Prescott, 
Arizona,  J.  F.  Blandy,  Amer.  Inst.  M.  E.,  XI,  291;  Timbering  in  Comstock 
Mines,  R.  W.  Raymond,  Amer.  Inst.  M.  E.,  VIII,  91;  Mine  Timbering  in 
the  Old  Ironsides  and  Knob  Hill  Mines,  H.  P.  DePencier,  Can.  Min.  Rev., 
July  31,  1902;  Drift  Timbering,  W.  H.  Storms,  Min.  &  Sci.  Press,  Oct.  10, 

i9°3- 

Timbering  in  German  Collieries,  Coll.  Guard.,  Oct.  24,  1902;  Coal-mine 
Timbering,  H.  W.  Halbaum,  Coll.  Eng.,  Feb.  1897,  303;  Drawing  of  Timber, 
E.  B.  Wain,  M.  &  M.,  XVIII,  159;  Colliery  Timber  and  Timber  Calculations, 
George  Johnson,  Coll.  Guard.,  Dec.  8,  1899;  The  Different  Methods  of  Sup- 
porting the  Roof  and  Sides  and  of  Resisting  the  Thrust  of  the  Floor  in  Coal 
Mines,  a  Prize  Essay,  Wm.  Bradford,  Coll.  Mgr.,  1896,  564;  Timbering 
Props  in  Coal  Mines,  Prize  Essay,  Coll.  Mgr.,  Nov.  20,  1896,  564;  Falls  of 
Roof  and  Sides,  Accidents,  Prevention,  Mine  Insp.  Coll.  Mgr.,  1894,  22. 

Creosoting  of  Timber  Sables  d'Olonne,  Coll.  Guard.,  LXXI,  886;  Wood 
Preservation,  H.  Flad,  Bull.  No.  i,  U.  S.  Department  of  Agriculture,  66; 
Behavior  and  Decay  of  Wood,  Bull.  No.  i,  U.  S.  Department  of  Agriculture; 
The  Best  Woods  for  Mines,  B.  E.  Fernow,  Am.  Inst.  M.  E.,  XVII,  269;  The 
Timber-preserving  Plant  of  the  Great  Northern  R.  R.,  Gaz.,  May  30,  1002; 
Timbering  Supporting  Roof  and  Sides  (Prize  Essay),  Coll.  Mgr.,  Nov.  20, 1896, 
564;  Timbering:  Economical  Use  of  Timber  in  Mines,  Coll.  Guard.,  Aug. 
28,  1896,  407;  Preservation  Method,  Calif.  Bureau,  i3th  Report;  Preserva- 
tion of  Timber,  Calif.  Mineralogist,  i3th  Report,  1896,  3d  Biennial,  647; 
Methods  of  Preservation,  Illustrated,  and  Table  of  Solutions  Used,  Amts., 
etc.,  The  Transit,  I,  63;  Treatment  of  Timber  for  Mine  Use,  R.  Martin, 
M.  E.,  X. 

The  Standard  Chute  Discharge  for  Mill  Holes,  M.  &  M.,  Vol.  XXIV, 

254- 


5?6  MANUAL  OF  MINING. 

Steel  Shaft-lining,  M.  &  M.,  Vol.  XXIII,  128,  155;  The  Use  of  Iron  Sup- 
ports in  Main  Roads  of  Mines,  instead  of  Masonry  or  Timbering,  G.  Meyer, 
Vol.  XXXVII,  221. 

Systematic  Timbering  in  French  Mines,  Coll.  Guard.,  Vol.  LXXXIV, 
1174;  Setting  of  Pit  Props,  Coll.  Guard.,  Vol.  LXXIX,  500;  Systematic 
Timbering,  Coll.  Guard.,  Vol.  LXXIX,  1075;  Erecting  Dams,  Coll.  Guard., 
Vol.  LXXXI,  569;  Timbering  German  Mines,  Coll.  Guard.,  Vol.  LXXXI, 
693;  Cribs  in  Collieries,  Coll.  Guard.,  Vol.  LXXXI,  570. 


CHAPTER  IV. 

DRIVING  DRIFTS,  TUNNELS,  AND  GANGWAYS. 

Levels. — These  horizontal  openings  in  metal-mines  receive 
an  amount  of  care  commensurate  with  their  importance  and 
service.  Like  the  levels,  which  also  are  driven  through  hard 
rock,  they  are  classed  as  "dead  work."  The  difficulties  of  driving 
are  not  so  great  as  for  shafts,  as  their  area  is  smaller  and  seepage 
water  gives  less  annoyance.  Their  location  is  fixed  by  condi- 
tions discussed  in  Chapter  I.  The  drifts  and  cross-cuts  of  a 
given  mine  are  of  the  same  dimensions.  They  serve  for  one  track, 
and  only  at  the  landings  are  wider  to  accommodate  a  second 
track.  They  are  rarely  timbered,  for  the  country  rock  will  stand 
without  support  for  the  entire  period  of  their  utility.  If  they 
require  protection  at  all,  a  masonry  or  iron  lining  is  used. 

Adits  and  'Levels  in  the  lode  are  secured  as  shown  in  the 
previous  chapter.  Usually  the  stull  and  lagging  suffice  for  the 
average  length  of  time  the  drift  is  kept  open.  Gangways  and 
galleries  in  coal-mines  have  a  larger  area  than  levels,  and  being 
exposed  to  more  treacherous  conditions,  possess  better  examples 
of  the  timberman's  art.  Tunnels  for  railway  or  drainage  pur- 
poses must  be  walled. 

Before  locating  a  tunnel  of  any  importance  a  careful  study 
of  the  ground  is  requisite.  All  the  data  obtainable  from  geolog- 
ical reports,  borings,  etc.,  should  be  availed  of.  The  character 
of  the  strata,  their  pitch,  and  the  direction  of  the  subterranean: 
drainage  should  be  known  before  the  character  of  timbering  and 
the  dimensions  and  amount  of  timber  required  can  be  determined- 

577 


S7»  MANUAL  OF  MINING. 

In  Drinker's  "Tunnelling"  will  be  found  a  discussion  of  the 
geological  conditions  affecting  tunnel  locations. 

The  Alignment  of  a  Tunnel. — The  axis  of  a  tunnel  is  kept  in 
a  vertical  plane  by  the  use  of  three  plumb-bobs  (Fig.  261),  and 
its  horizontaiity  is  tested  by  a  long  spirit-level.  A  grade  of  two 
feet  in  a  hundred  is  given  to  the  floor  for  purposes  of  drainage.  In 
collieries  the  grade  is  involved  with  the  proposed  system  of  mining. 

The  alignment  of  a  gentle  incline  is  similarly  conducted, 
though  four  adjustments  are  necessary.  A  saw-cut  is  made 
across  each  sill  and  8  inches  from  the  end.  Two  hubs  are  set 
in  the  sills  on  this  line  by  transit  within  50  feet  of  the  face  of  the 
drift  or  incline.  They  are  lined  by  a  string  which  must  pass 
through  the  saw-cut  notches.  The  grade  stick  is  now  placed  in 
position  to  depress  or  raise  the  sills  to  the  required  inclination, 
after  which  they  must  be  laid  level  and  across  the  drift  at  right 
angles  with  the  line. 


FIG.  261. — Aligning  a  Tunnel  Drift. 

Rock  Characteristics. — Hard  rock  presents  no  serious  dif- 
ficulties to  the  miner  beyond  the  time  and  cost  of  drilling,  for 
it  usually  affords  a  good  roof  and  if  uncre viced  is  dry.  Granite 
dolomite  and  gneiss  are  of  this  character.  Slates  and  shales 
are  bad  and  require  arching  or  block  timbering.  Porphyry  is 
treacherous,  because,  though  hard  when  first  opened  into,  it 
soon  decrepitates  on  exposure.  Some  of  the  limestones  and 
sandstones  are  porous  and  wet.  Clay  seams  are  bad  primarily, 
besides  being  watercourses  for  the  upper  porous  strata. 

Driving  in  Creviced  Rock. — In  uncreviced  rock  the  order 
of  the  breaking  is  immaterial,  so  long  as  a  good  bench  is  had 
to  shoot  from  and  a  favorable  working-face  is  obtained  for  the 


DRIVING  DRIFTS,    TUNNELS,   AND  GANGWAYS. 


579 


next  shot.  The  crevices  and  planesi  n  stratified  ground  materi- 
ally assist  the  shooting.  If  the  seams  pitch  toward  the  face 
(Fig.  262),  the  upper  holes  are  fired  first,  or  in  high  galleries 
the  central  holes  will  make  a  good  bench  to  shoot  to  from  top 
to  bottom.  If  the  cleavage  is  not  marked,  or  away  from  the 
face  (Fig.  263),  the  bottom  holes  are  fired  before  the  upper  ones. 
If  electric  firing  is  adopted,  the  sequence  of  shots  is  of  little  im- 
portance. Three  men  form  a  gang  in  driving,  and  can  manage 
the  ordinary-sized  drift.  In  the  gangway  of  7X7  two  machines 
or  two  pairs  of  cutters  have  ample  room.  Their  progress  cannot 

FIG.  262 


Placing  Holes  in  Creviced  ROCK. 

generally  be  stated.  In  hard  rock  a  foot  a  shift  is  fair,  while 
soft  rock  can  be  penetrated  at  the  rate  of  three  feet  a  day.  In 
shaly  ground  greater  progress  might  be  made  did  it  not  require 
good  timbering.  In  soft  ground  the  advance  depends  upon 
the  skill  of  the  timberman. 

Driving  by  Air-drills. — It  is  a  matter  of  indifference  where  the 
holes  are  placed  for  explosives  if  machine-drills  and  simultane- 
eous  firing  are  employed.  A  drift  of  ordinary  size  will  easily 
accommodate  one  drill,  wbile  two  can  advantageously  work  in 
a  heading  10  feet  wide,  obtain  more  angling  holes,  and  advance 
more  rapidly.  In  railroad  tunnels  four  are  simultaneously  drilling. 
Hand-work  is  much  more  depended  upon  for  the  driving  of 
gangways,  but  machine-drills  are  becoming  popular  for  tunnels. 
Undoubtedly  the  progress  by  machine  is  greater  than  by  hand, 
while  the  cost  is  nearly  the  same  per  lineal  foot.  Where  the  drills 
and  air-compressors  can  be  put  to  use  after  the  mine  i?  opened, 


580  MANUAL  OF  MINING. 

it  certainly  would  be  advantageous  to  employ  them  in  the  pre- 
liminary and  development  work. 

The  consumption  of  steel  in  medium  ground  is  about  25 
cents  per  cubic  yard  removed,  though  pink  quartz  will  dull  150 
bits  to  the  hole,  and  the  blacksmith  will  consume  more  steel 
than  the  rock.  The  consumption  of  powder  is  about  $i  per 
cubic  yard  of  medium-tough  rock  removed.  This  amount  will 
ry  with  the  area  of  the  face  and  whether  the  breaking  is  done 

simultaneous  or  single  shots. 

Driving  Slopes. — The  operations  of  sinking  and  the  timbering 
of  a  slope  are  similar  to  those  of  level  driving,  except  that  the 
sill  is  indispensable  to  the  set.  It  must  be  well  bedded  and 
let  into  the  rocks  on  each  side  to  prevent  the  roadway  from  slip- 
ping down  hill.  Ofttimes  it  is  stayed  by  plugs  driven  into  the 
floor.  The  sets  are  also  braced  against  one  another  by  longitu- 
dinal studs  between  the  posts  at  their  head. 

Driving  Tunnels. — Levels  and  drifts  are  worked  over  their 
entire  face.  But  when  tunnels  are  to  be  driven  of  a  height  greater 
than  8  feet,  they  must  be  broken  in  benches.  Railroad  tunnels  are 
of  this  description,  and  in  hard  rock  may  be  driven  without  any 
temporary  timbering,  the  benches  being  attacked  like  stopes 
(Fig.  27),  where  the  numbers  express  the  sequence  of  openings. 
Frequently,  in  driving  long  tunnels,  the  drift  No.  i  is  pushed  as 
fast  as  possible  in  order  to  make  connection  with  a  shaft  or  a 
similar  drift  approaching  from  the  opposite  direction.  This  com- 
munication is  for  ventilation  and  haulage  purposes. 

Numerous  tunnels  have  been  driven  more  than  five  miles 
for  mining  purposes.  The  Freiberg  is  24  miles  long;  at  Claus- 
thal  is  one  nearly  n;  the  Joseph  II.  is  9^;  the  Ernest  August, 
6£  miles;  and  the  Sutro,  5.  These  have,  moreover,  lateral 
branches  that  enable  them  to  subserve  a  great  territory.  The 
Gwennap  adit,  in  Cornwall,  is  said,  with  its  branches,  to  attain 
a  length  of  30  miles.  The  greatest  depth  below  the  surface  is 
not  over  900  feet  for  any  one  of  these. 

The  Use  of  Supplementary  Shafts. — Auxiliary  shafts  are  sunk 
at  convenient  points  along  the  line  of  a  long  tunnel  to  its  level; 


DRIVING  DRIFTS,   TUNNELS,  AND   GANGWAYS.          58* 

and  from  them  the  excavations  are  begun  contemporaneously 
with  those  at  the  mouths  of  the  tunnel.  Sometimes  the  shafts 
are  to  one  side  of  the  tunnel  line,  to  keep  them  free  and  clear 
of  the  tunnel-work.  The  amount  of  time  allowed  for  the  com- 
pletion of  the  tunnel  determines  the  number  of  points  of  attack- 
and  the  number  of  shafts  to  be  sunk.  The  latter  is  also  depend- 
ent upon  the  cost  of  sinking  and  the  hoisting  through  them  rela- 
tive to  that  of  the  long  haul  in  the  tunnel.  These  relations  car* 
be  mathematically  expressed  when  the  several  aspects  of  the  case- 
are  given.  From  Foster's  Gallon's  "Lectures  on  Mining"  the 
following  formula  is  taken: 


wherein  17  =  the  total  known  cost  of  shaft,  x=  number  of  shafts 
to  be  opened  over  a  length  I,  5  =  the  area  excavated,  S'  —  the 
section  of  the  lining,  Z)  =  the  distance  between  the  two  adjoining 
shafts,  P  and  P'  =  the  cost  of  haulage  per  lineal  yard  for  each 
cubic  yard  of  rubbish  and  of  walling  material. 


The  St.  Gothard  Tunnel,  48,840  feet  long,  was  run  from  the 
two  ends  only;  so  the  Mont  Cenis  Tunnel,  39,840;  the  Hoosac 
and  the  Sutro  tunnels  with  the  assistance  of  two  shafts.  The 
Washington  Tunnel,  20,715  feet,  had  four  working  shafts.  The 
Rothschenberger  mining  adit  had  18  points  of  attack  along  its 
24  miles  of  length.  Rziha  estimates  that  the  "additional  cost 
of  running  headings  from  a  shaft  is  from  5  to  10  per  cent  higher 
than  running  from  portals."  He  also  gives  a  unique  table  to 
show  that  the  rate  of  progress  per  month  is  apt  to  be  greater  in 
long  than  in  short  tunnels.  Of  those  less  than  100  m.  long, 
29  feet  advance  was  the  average  of  3;  13  between  400  and  600  m. 
showed  average  of  57;  8,  up  to  1000  m.,  114  feet,  while  6  between 
3000  and  4000  m.  had  an  average  of  219  feet,  and  4  over  4000  m, 
progressed  259  feet  per  month. 


58' 


MANUAL   OF  MINING. 


Provision  for  traffic  is  not  a  simple  matter  when  it  is  recalled 
that,  besides  the  actual  mining  of  an  area  (say,  for  example, 
that  of  the  St.  Gothard  Tunnel),  26X20,  at  a  rate  of  18  feet  per 
day,  and  the  placing  of  1000  cu.  ft.  of  timber  and  300  tons  of 
masonry  per  day,  750  tons  of  broken  rock,  15  tons  of  lumber, 
•and  300  tons  of  masonry  must  be  handled,  loaded,  transported, 
.and  unloaded  at  the  same  time. 

As  the  hoisting  through  shafts  is  about  half  as  fast  as  haul- 
age through  a  heading,  the  use  of  auxiliary  shafts  is  not  so  eco- 
nomical as  it  was  before  the  present  development  of  machine- 
drilling  and  blasting  methods. 

Driving  a  Tunnel. — There  are  several  methods  of  opening 
a  tunnel,  one  of  which  is  depicted  in  Fig.  264.  The  two  side 


FTG.  264. — The  Order  of  Attack  on  a  Tunnel-face. 

•drifts  at  the  bottom  may  precede  the  main  excavation;  in  these 
are  built  the  walls  for  the  roof  arch  which  ultimately  lines  the 
tunnel.  While  the  small  drift,  i,  is  being  driven  to  the  connec- 
tion, the  sides  2,  2  are  following,  and  in  firm  ground  they  may 
precede  5,  5  at  the  bottom  by  150  feet.  At  this  point  the  tunnel 
is  being  lined  with  timber  or  masonry,  while  gangs  are  also  break- 
ing ground  on  benches  3  and  4. 

If  the  roof  needs  support,  the  plan  of  work  contemplates 
one  similar  to  that  of  Fig.  265,  with  the  exception  that  the  central 


DRIVING  DRIFTS,  TUNNELS,  AND  GANGWAYS.          583 

core  3,  4  remains  as  a  support  to  the  timbers  arranged  as  shown 
below.  Inside  of  this  frame  the  masonry  may  be  completed, 
after  which  the  inverted  arch,  if  required. 

In  soft  ground,  and  especially  in  inclined  strata,  the  order 
of  driving  the  benches  and  the  amount  and  character  of  tim- 
bering varies  according  to  the  dip.  Circumstances  and  diffi- 
culties are  so  diversified  that  no  uniform  infallible  rule  has  been 
established  for  the  guidance  of  the  engineer — probably  because 
no  system  has  a  superiority  over  all  others  for  any  and  all  con- 
ditions. Safety  is  an  clement  as  important  as  time  and  money> 
and  often  is  a  determinant. 


FIG.  265. 

Tunnelling  Systems. — A  brief  review  of  the  English,  Belgian,, 
German,  and  Austrian  is  taken  from  Drinker's  ''Tunnelling." 

The  English  system  is  developed  from  the  experiences  in 
the  Thames  River  tunnel  and  consists  in  taking  out  the  full  area 
at  once,  after  the  preliminary  top-heading  has  been  made,  and 
in  supporting  the  roof  by  longitudinal  top  bars  while  removing 
the  lower  section.  This  gives  a  full  clear  area  for  putting  in. 
the  maSonry,  after  which,  in  material  having  any  tendency  to 


584  MANUAL  OF  MINING. 

swell,  the  space  behind  the  side  walls  is  securely  grouted.  Though 
built  quicker  and  applicable  in  90  per  cent  of  cases,  it  is  unsatis- 
factory in  very  bad  ground;  in  heavy  ground  it  requires  more 
limber  than  does  the  Austrian,  its  strongest  competitor. 

The  Belgian  method,  introduced  after  the  iron  shield  had 
t>een  tried  ineffectually  in  quicksand,  builds  the  tunnel  as  an 
open  cut  down  to  the  springing-line  of  the  arch.  The  arch  is 
ihen  laid,  recovered,  and  underpinned,  the  bottom  removed  in 
benches,  working  downwards,  and  the  abutments  built.  Some- 
times the  centre-core  is  left,  though  that  is  only  done  in.  the 
Trench  and  German  modifications.  The  entire  area  is  not 
attacked  at  once,  but  divided  into  several  benches,  each  being 
worked  separately.  The  underpinning  of  the  arch  may  be  safe 
enough  in  hard  ground,  but  it  certainly  affords  a  doubtful  security 
in  loose  material. 

The  German  system  gave  rise  to  what  may  be  called  the 
centre-core  system.  The  work  of  excavation  begins  at  the  foot 
of  each  abutment  of  the  arch,  where  a  small  heading  (Fig.  265) 
with  timber  sets  is  first  driven.  In  this  the  foundation  is  laid; 
above  it  a  second  heading,  large  enough  to  build  another  height 
of  wall;  above  this  another,  in  which  the  masonry  is  carried 
up.  The  top  is  then  excavated  across,  and  a  connection  effected 
between  the  two  sides.  In  this  the  arch  is  completed  without  the 
use  of  centres,  while  the  roof  is  being  supported  by  stulls  or  props. 
The  core  is  then  removed. 

In  hard  ground  it  gives  cheap  working,  from  the  fact  that 
the  core  has  several  faces  of  attack.  In  soft  ground  it  is  safer, 
because  of  the  small  exposure  of  roof  and  face;  and  the  centre- 
core  saves  timber.  Its  ventilation  is  bad,  and  the  cost  of  lay- 
ing masonry  is  larger  than  where  the  masons  have  elbow-room; 
it  is  hard  to  securely  timber,  and  several  prominent  engineers 
have  decided  against  it.  Certainly,  in  soft,  treacherous  ground, 
like  shales  and  clays,  its  defects  disqualify  it. 

The  Austrian  system  admits  of  mining  the  whole  area  in  small 
sections.  First  a  bottom  heading  is  driven  and  afterwards  con- 
.nected  to  a  top  heading,  which  is  finally  widened  to  full  width 


DRIVING  DRIFTS,  TUNNELS,  AND   GANGWAYS.          585 

for  a  bar-timbering  which  is  carried  down  to  the  springing-line. 
Sides  are  excavated  for  the  walls,  after  which  foundations  are 
prepared  and  the  side  walls  built;  finally,  the  invert  arch.  The 
cross-rafter  timbering  (Figs.  267  and  296)  used  for  support  ad- 
mits of  the  transfer  of  pressure,  and  there  is  no  such  undue  con- 
centration as  in  Fig.  265.  This  disposition  of  the  timber  affords 


FIG.  266. 

a  greater  strength,  and  is  the  leading  feature.  Additional  braces 
are  sometimes  added  and  the  sets  connected  every  which  way; 
but  the  design  is  to  arrange  the  timber  of  each  section  in  such 
manner  as  to  form  an  integral  part  of  the  completed  system. 
After  the  roof-timbers  are  in  place,  plenty  of  space  is  left  below 
for  masons,  and  good  ventilation  is  had.  « 


586  MANUAL  OF  MINING. 

Comparison  of  the  Four  Systems. — Each  system  has  dis- 
tinctive features  commending  it  to  favor  under  given  conditions, 
and  a  brief  comparison  may  be  given. 

The  centre-core  system  is  emphatically  inapplicable  for 
loose  rock  where  other  methods  are  better.  It  is  quite  expen- 
sive in  all  of  its  operations,  and  is  not  safe. 

The  English  system  is  safe  enough  for  most  tunnels  that 
are  likely  to  be  driven,  and  offers  ample  space  for  the  miners 
and  masons  to  work,  but  it  is  not  adapted  to  very  bad  ground 
or  where  heavy  pressure  is  encountered. 

The  Belgian  system  seems  to  be  preferred  for  moderately 
firm  ground,  in  which  it  is  quicker  than  either  of  the  two  above 
mentioned;  but  it  is  hardly  suited  to  running  ground,  because 
of  the  great  care  which  must  be  taken  with  the  underpinning. 
The  cost  of  transportation  while  building  is  higher  than  in  the 
others  because  of  the  frequent  interruptions  caused  by  the  trans- 
portation scaffolding.  In  very  loose  ground  it  is  positively 
dangerous. 

The  Austrian  system  has  more  to  commend  it  and  less  to 
condemn  it  than  the  others.  All  the  driving  operations  are 
cheaply  and  easily  conducted.  The  driving  is  least  interrupted, 
and  safety  is  assured  if  the  middle  timbering  and  the  longitu- 
dinal bracing  be  well  apportioned.  The  forepoling,  if  em- 
ployed, would  be  in  the  direction  of  the  tunnel  axis,  and  the 
masonry  arching  is  not  delayed. 

Block  Timbering. — In  Figs.  266  and  268  are  illustrated  the 
timbering  permanently  placed  in  the  tunnel.  Each  of  the  nine 
blocks  is  io"Xio",  the  legs  being  12  inches  square.  From  the 
three- rafter  system  of  roofing  the  tunnel  American  engineers 
developed  the  block-timber  arching  for  the  support  of  the  roof 
rock  which  is  very  loose  and  requires  close  bracing.  The  wooden 
voussoirs  were  first  used  for  temporary  support  only,  until  they 
were  replaced  by  brick  or  stone  arching,  but  latterly  they  have 
served  as  a  permanent  arching  lined  with  a  fire-proof  sheeting 
for  railroad  use. 

We  cannot  claim  any  system  of  tunnelling  as  our  own,  for 


DRIVING  DRIFTS,   TUNNELS,  AND   GANGWAYS. 


587 


neither  the  number  of  tunnels  nor  the  difficulties  encountered 
are  as  great  as  in  the  Old  World.  The  Austrian  method  is  the 
nearest  approach  to  ours,  or,  rather,  is  the  one  which  our  en- 
gineers have  adopted  with  a  modified  framing.  The  mode  of 
driving  and  timbering  is  illustrated  in  Fig.  264.  The  upper 
heading,  i,  with  its  enlargement,  2,  2,  precedes  the  work  upon 
the  "bench."  With  a  tunnel  area  of  2i"X27"  the  heading  is 
about  8  feet  high,  and  the  bench,  the  remainder,  attacked  in 
one  section.  The  upper  heading  is  timbered,  rafter  style,  three 
or  nine  voussoir  bents  forming  a  block-arch  on  heavy  perma- 
nent pole  standards,  or  on  the  side  rock  if  sufficiently  sound.  In 
loose  rock  the  number  of  bents  increase  and  the  timbering  is 
heavier.  Inside  of  this  frame  arch  the  masonry  is  built.  A 
segmental  arch  21  feet  high  by  18  wide  in  the  clear  had  nine 
voussoir-blocks  25  inches  long  and  io"Xio"  section;  two 
wall-plates,  6  feet  4  inches  long,  12  inches  square;  two  posts, 
14'  i"Xio"Xi2";  and  45  lagging,  6'  6"X6". 

Fig.  268  illustrates  the  style  of  bar-timbering,  which,  how- 
ever, will  not  permanently  withstand  much  lateral  pressure. 


FIGS.  267  and  268  —The  Block  Arching  of  Tunnels. 

Tunnelling  Soft  Ground. — In  running  ground  the  area  at- 
tacked is  limited  by  the  rapidity  with  which  permanent  support 
can  follow  the  excavation,  during  which  provision  must  also  be 
made  against  the  pressure  coming  from  all  sides.  In  no  region, 
not  even  in  the  Rocky  Mountains,  is  the  engineer  free  from  the 
liability  of  striking  ground  that  may  overwhelm  the  miner  before 
timbers  can  be  inserted;  so  that  the  ground  has  to  be  restrained 
up  to  the  face  as  well  as  behind  the  frames,  with  which  close 


S88 


MANUAL  OF  MINING. 


lagging  may  suffice  to  prevent  movement.  If  it  is  not  in  new 
ground,  the  consistency  of  which  will  determine  the  details  of 
its  penetration,  it  may  be  where  timbers  have  rotted  or  given 
way  that  a  cave  or  a  run  may  occur. 

In  alluvial  or  sand  the  method  of  spilling  or  a  pneumatic 
system  is  indispensable.  These  have  been  employed  under  the 
waters  of  the  Thames,  the  Severn,  and  the  Mersey  Rivers  in 
Great  Britain  and  are  universally  employed  for  driving  through 
similar  ground  in  America. 

Spilling  by  laths  is  a  method  equally  applicable  to  drifts  as 
to  shafts.  In  front  of  one  stick  of  a  set,  and  behind  its  mate  in  the 
next  advancing  set,  pointed  heavy  planks 
are  driven,  one  set  in  advance  of  the 
face,  close  together  on  one  or  all  of  the 
sides  from  which  the  pressure  is  exerted. 
The  fore  end  of  the  plank  is  forced 
down  upon  its  set,  while  the  rear  end 
is  held  against  the  lower  side  of  its  cap, 
being  protected  from  pressure  by  the 
previously  driven  upper  lath.  This  is  the 
method  of  forepoling  illustrated  in  Fig.  269 
The  progress  in  soft,  heavy  timbered  rock 
is  three  times  as  fast  as  in  very  solid  rock. 
The  "spilling"  protects  the  fore- 
breast  by  horizontal  laths,  as  long  as  the 
breast  is  wide,  held  against  it  and  braced 
to  the  nearest  set.  Each  lath,  a  (Fig. 
270),  is  removed  in  descending  order  to 
permit  some  ground  to  run  off;  it  is 

advanced  a  short  distance  and  braced  again,  b.  The  progress 
depends  upon  the  speed  with  which  the  spaces  may  be  opened 
and  closed;  as  these  are  small,  the  movement  is  controllable, 
and  there  need  be  no  fear  of  sudden  shock  to  the  timbers.  The 
method  is  simple,  and  has  been  eminently  successful  under  many 
circumstances;  once,  with  masonry  lining  following  the  fore- 
pole,  a  tunnel  was  executed  within  18  feet  of  a  river-bed. 


I  it  HUM 

FIG.  269. — Forepoling. 


DRIVING    DRIFTS,  TUNNELS,  AND   GANGWAYS.         589 


FIG.  270. — Method  of  Spilling  for  Running  Ground. 


59° 


MANUAL  OF  MINING. 


An  entirely  different  principle  is  that  employed  by  Durieux 
and  others  in  Westphalia,  whereby  the  ground  was  forced  ahead 
of  the  work  and  was  not  removed  at  all  (Fig.  271).  The  walls 


FIG.  271. — A  Picketing  Method  for  Sandy  Ground. 

and  roof  were  forepoled,  but  the  breast  and  floor  were  checkered 
by  pyramidal  pickets,  completely  covering  the  exposure.  Those 
on  the  face  were  square  faced  and  larger  than  the  floor-wedges, 
which  were  12  inches  long  and  4  inches  diameter,  driven  by 
mallets.  The  floor-pickets  remain,  in  permanently,  while  those 
at  the  face  force  the  soft  material  ahead  and  advance  in  this 
manner.  The  battering  they  receive  renders  them  useless  in  a 
very  short  while,  when  they  are  replaced.  There  is  no  material 
to  be  hauled  except  that  used  in  the  construction,  and  rather 
bad  ground  has  been  thus  traversed.  On  occasion  a  short 
lateral  drift  is  similarly  pushed  to  relieve  the  main  work.  Four 
men  in  a  drift  s'X6'  in  morainal  matter  will  advance  4  feet  a 
shift. 

The  finest  piece  of  tunnelling  is  the  construction  of  the  new 
Croton  Aqueduct,  where  water,  mud,  quicksand,  and  all  varie- 
ties of  loose  soil  were  penetrated  in  an  area  of  676  sq.  ft.  Fore- 
poling  for  the  roof  and  sides,  picket- spill  ing  for  the  floor,  and 
the  American  system  with  block-arching  were  adopted  at  various 


DRIVING  DRIFTS,  TUNNELS,  AND   GANGWAYS.          59 1 


FIG.  272. — The  Brunei  Shield. 


592  MANUAL  OF  MINING. 

points.  The  ground  in  extended  places  was  so  bad  that  24- 
inch  timbers  were  crushed  by  the  pressure. 

Iron  Shields  at  the  Tunnel  Face. — Iron  is  a  safer  constructive 
material  for  loose  soil;  and  in  1825  Brunei  put  it  to  use  for  exca- 
vating an  area  38X22  under  the  Thames  River.  As  in  the  spill- 
ing method,  the  face  was  covered  by  laths  about  3"X3"X6", 
and  for  rapidity  of  work  was  subdivided  into  three  tiers  of  n 
breasts  each  (Fig.  272),  each  being  protected  by  a  cast-iron  shield, 
from  which  struts  hold  the  laths  d  in  place.  One  man  to  a  cell 
operates  by  taking  away  a  lath  and  replacing  it  3  inches  in  ad- 
vance. The  laths  are  removed  successively  downwards  until  3 
inches  of  progress  has  been  made,  after  which  the  alternate  frames 
are  carried  ahead  6  inches  and  the  performance  repeated.  In 
each  cell  is  a  similar  performance.  The  masons  follow  the  miners 
very  closely  at  G. 

The  second  Thames  Tunnel  was  completed  by  the  use  of  a 
shield,  which  with  pointed  shoes  was  forced  into  a  stiff  clay  at 
the  rate  of  9  feet  per  day  by  jack-screws  exerting  a  pressure  of 
60  tons  on  it.  Three  men  crawled  through  a  door  in  its  face 
and  excavated  some  earth  preparatory  to  the  next  move. 

The  Hydraulic  Shield. — The  present  practice  for  subaqueous 
tunnels  involves  some  variation  of  the  pneumatic  system.  A 
pilot-tube  or  caisson  penetrates  the  soil,  which  is  held  up  by 
compressed  air.  The  masonry,  perforce,  is  built  as  fast  as  the 
shield,  tube,  or  caisson  advances.  In  the  Greathead  system  a 
cylindrical  iron  shield,  21  feet  in  diameter,  is  thrust  from  the 
masonry  by  hydraulic  pumps  under  a  pressure  of  3000  Ibs.  per 
square  inch.  The  front  edge  of  the  shield  is  a  heavy  ring  shoe 
to  facilitate  its  advance  into  the  silt.  The  rear  end  of  the  shield 
encloses  the  masonry,  which  follows  the  progress  of  the  shield. 
Doors  in  its  face  are  opened  to  allow  the  soft  earth  to  enter. 
This  is  promptly  removed  (Fig.  273). 

In  preparing  to  tunnel  silt,  both  the  weight  and  the  vertical 
pressure  of  the  overlying  material  and  the  lateral  movement 
of  the  loose  paste  are  to  be  resisted.  The  first  is  a  matter  of 
determination,  and  the  ability  of  the  completed  structure  to 


DRIVING  DRIFTS,  TUNNELS,  AND  GANGWAYS.          593 


FIG.  273.— The  Hydraulic  Shield, 


594 


MANUAL  OF  MINING. 


withstand  this  is  also  a  matter  of  mathematical  calculation; 
but  the  second  is  the  difficulty  to  be  apprehended.  The  Hudson 
River  Tunnel  engineers,  however,  satisfied  themselves  that  the 
tendency  of  the  gravel  to  pour  over  the  top  of  the  tube  would 
reduce  the  lateral  stress  to  the  resistance  of  the  tube.  Certainly 
their  experience  in  subaqueous  driving  goes  to  show  that  the 
tube  serves  better  than  the  timbering  systems  as  regards  the  pre- 
vention of  overhead  settlement.  The  liability  to  settlement  in 
front  of  the  shield  may  be  overcome  by  grouting  under  pressure 
at  the  rear  of  the  shield  between  the  lining  and  the  roof. 

The  Anderson  Pilot-tube. — This  is  a  segmental  tube  of  about 
6  feet  diameter,  strengthened  inside  by  radial  timbers,  which 
precede  the  main  work.  It  is  of  -J-inch  plates,  I2"X24",  riveted 
together  by  means  of  flanges;  and  when  a  cut  has  been  excavated 
into  the  heading  large  enough,  one  of  the  plates  is  placed  and 


RTClN^L^^ 


FIG.  274. — The  Anderson  Pilot -tube. 

held  by  props  (often  the  plates  arc  held  by  compressed  air  during 
the  work) ;  on  each  side  other  cuts  are  made  for  two  more  plates, 
which  are  riveted  to  it.  Rings  of  the  pilot  are  thus  successively 
completed. 

Around  this,  in  small  terraces,  and  considerably  behind  the 
pilot,  the  main  shell,  17  feet  in  diameter,  is  finished  in  a  similar 
manner,  the  plates  being  propped  from  the  pilot-tube,  which  is 
always  braced  from  the  masonry'  that  lines  the  shell.  With  its 
progress  the  rear  rings  of  the  pilot-tube  are  removed  and  their 
plates  shifted  to  the  front  end.  The  masonry  consists  of  six 
courses  of  brick  laid  in  cement.  To  reduce  the  volume  of  the 


DRIVING  DRIFTS,  TUNNELS,  AND   GANGWAYS.          595 

tunnel  that  is  kept  under  the  compressed  air,  brick  bulkheads, 
4  feet  thick,  provided  with  two  air-locks,  are  built  every  400 
or  500  feet.     Only  the  two  nearest  the  work  are  maintained. 
The  following  references  are  cited: 

REFERENCES. 

Tunnel  Shields  for  Driving,  Eng.  News,  Jan.  9,  1892,  26;  Shields,  Eng. 
Record,  Mar.  25,  1893,  337;  Lining,  Jour.  Assn.  Eng.  Soc.,  July  1893,  331; 
Lining,  Eng.  News,  Oct.  12,  1893,  289;  Tunnelling  with  Steam  Shovels, 
M.  &  M.,  Vol.  XX,  128;  Alignment  of  Tunnels,  E.  R.  Durham,  Am.  Soc. 
C.  E.,  Oct.  12,  1893,  289;  Methods  in  Tunnelling  in  Soft  Grounds,  W. 
Beaham,  A.  Soc.  C.  E.,  XXVI,  490;  Lining  Tunnels,  E.  R.  McNeill,  Engi- 
neering News,  Oct.  12,  1893,  289;  Timber  Lining  of  Tunnels,  S.  B.  Fisher, 
Engineering  News,  XXX,  268;  Lining  Galleries  Exposed  to  Heavy  Rock 
Pressure,  Bergassessor  Jacob,  Coll.  Guard.,  Dec.  12,  1902. 

The  Ventilation  of  the  St.  Gothard  Tunnel,  Rev.  Gdn.  des  Chemins  de 
Fer,  Nov.  1899;  The  Cripple  Creek,  Colo.,  Drainage  Tunnel,  W.  B.  Wilson, 
Min.  &  Sci.  Press,  Jan.  17,  1903;  The  Ventilation  of  the  Simplon  Tunnel,  C. 
I.  Wagner,  Oesterr.  Wochenschr.  f.  d.  Oeffent.  Baudienst,  Feb.  15,  1902; 
Methods  of  Driving  Tunnels  in  Soft  Ground,  Anon.  Eng.  News,  Jan.  16, 
1892,  64;  Driving  Shields,  M.  P.  Paret,  Eng.  News,  Jan.  9,  1892,  26;  Tim- 
ber Lining  of  Tunnels,  S.  B.  Fisher,  XXX,  268;  and  The  Ventilation  of 
Tunnels,  Heating  and  Ventilating,  Serial,  1904. 


CHAPTER  V. 

DRILLING-  AND   BORING-MACHINES   FOR  EXPLORATIONS. 

Classes  of  Machines. — The  penetration  of  rock  or  of  loose 
material  may  be  accomplished  by  either  of  two  operations  repre- 
sented by  corresponding  types  of  machines.  By  a  repeated 
reciprocating  blow  or  concussion  upon  the  rock,  fragments  may 
be  broken  therefrom  and  a  hole  drilled  to  any  depth  and  at  any 
rate,  depending  upon  the  strength  of  the  machine  and  the  resist- 
ance of  the  rock  to  abrasion.  Machines  operating  upon  the 
rock  in  this  way  are  known  as  percussive  machines  or  punch- 
drills.  The  second  method  of  penetration  consists  in  employing 
a  rotary  cutter  which,  with  sufficient  pressure  or  weight  upon  it, 
will  abrade  the  rock  at  a  rate  depending  upon  its  relative  soft- 
ness and  the  motor  power. 

Percussive  drills  are  employed  to  advantage  upon  hard  uni- 
form rock  only.  Rotary  borers  are  constructed  of  various  types 
as  to  be  equally  applicable  to  either  soft  rock,  loose  ground,  or 
hard  rock. 

These  drills  or  borers  may  be  operated  by  directly  connected 
engines  upon  the  same  frame  to  drill  holes  of  thousands  of  feet 
in  depth  for  prospecting  purposes.  When  these  machines  are  to 
be  used  underground  they  are  of  a  more  portable  type  and  oper- 
ated by  electricity  or  compressed  air,  conveyed  to  it  from  a  dis- 
tinct source.  The  depth  of  the  holes  made  possible  with  these 
is  limited  to  a  few  feet.  Hence  their  utility  is  confined  to  drill- 
ing holes  for  blasting  purposes  and  not  for  exploration. 

596 


DRILLING-  AND  BORING-MACHINES.  597 

Prospecting-machines. — The  machines  employed  for  pros- 
pecting are  either  of  the  punch-drill  type  represented  by  the 
American  oil-well  rig,  or  the  rotary  types  represented  by  the 
"diamond"  drill  and  other  forms  of  tubular  drills. 

Bore-holes  are  driven  for  the  testing  of  the  strata;  to  prospect 
for  the  gas;  to  afford  an  outlet  for  water;  to  pump  brine  to  the 
surface;  or  to  subserve  some  precautionary  measures  in  mining, 
such  as  special  ventilating  openings,  deep  sump-holes  to  drain 
the  mine  into  a  lower  porous  stratum  or  conduits,  for  tail-haulage 
rope.  These  machines  are  also  employed  for  drilling  a  series  of 
holes  in  connection  with  the  Poetsch  method  and  the  long  hole 
process  for  shaft-sinking. 

Percussive  Drills. — Reciprocating  motion  is  obtained  from 
a  steam-  or  air-piston  either  directly  attached  to  the  drill  or  indi- 
rectly connected  with  it  by  rope.  The  blow  may  be  produced 
by  the  pressure  of  the  motor  fluid  in  the  first  case  or  by  the  weight 
of  the  drill  allowed  to  fall  freely  after  having  been  raised  by  the 
rope.  In  the  first  method  the  piston  pressure  is  great  compared 
with  the  weight  of  the  tool,  while  in  the  latter  the  force  of  the 
concussion  is  solely  due  to  the  heavy  weight  of  the  drill.  The 
early  type  of  punch-drill  consisted  of  a  continuous  line  of  iron 
rods,  lengthened  from  time  to  time  with  the  progress  of  the  hole, 
which  was  raised  and  lowered  at  the  surface  by  a  walking-beam 
operated  by  a  cam  or  other  lifting  device.  The  shock  transmitted 
through  the  long  line  of  metal  caused  breakage  of  the  tools,  and 
other  difficulties  which  led  to  the  later  forms  of  machines  in 
which  heavy  drill  lengths  were  suspended  from  the  surface  by  a 
flat  or  round  rope  and  at  a  much  more  rapid  rate  than  the 
line  of  rods.  Later  the  solid  rods  were  replaced  by  several  lengths 
of  hollow  rods  which  gave  greater  progress  with  the  same  power, 
and  overcame  the  difficulties  of  feeding  water  and  disposing  of 
the  slimes  produced  during  the  drilling.  The  rods  were  screwed 
together  or  bayonet-hooked,  the  former  being  preferred.  The 
engine  may  be  single-acting  with  a  reversible  link  motion  pro- 
vided with  a  sectional  fly-wheel,  whose  weight  was  increased  as 
the  hole  deepened.  It  communicates  motion  to  a  walking- 


598 


MANUAL  OF  MINING. 


beam  which  is  pivoted  at  its  centre  on  top  of  two  sampson- posts 
and  fastened  at  the  derrick  end  to  the  drill-rope  attachments. 

Oil-well  Rig. — The  typical  American  drill  rig  is  illustrated  in 
Fig-  275. 

This  consists  of  a  derrick  75  feet,  or  more,  in  height,  having 
at  its  crown  a  sheave  over  which  the  drill- rope  passes  to  the  "bull- 
wheel,"  or  large  drum.  A  high  derrick  is  required  to  facilitate 
the  attachment  and  the  removal  of  an 
entire  string  of  tools. 

A  second  rope  pulley,  at  one  side  of 
the  derrick  near  the  top,  guides  the  rope 
attached  to  the  sand-pump,  which  rope  is 
connected  with  its  reel.  The  several  bents 
of  the  derrick  are  fastened  together  by 
wedges,  if  of  .wood,  with  the  exception  of 
the  derrick  floor-planks,  which  are  bored 
or  spiked.  Its  construction  permits  of  easy 
dissection  when  a  hole  is  to  be  abandoned. 

The  Jars. — The  concussion  produced 
by  the  sudden  arrest  of  a  fall  of  a  mass  of 
iron  such  as  was  in  the  drill-rods  when 
transmitted  to  the  latter  caused  such  diffi- 


FIG.  275. — An  American  Drill  Rig. 

culties  as  were  sough  to  be  removed  by  various  appliances. 
Oennhausen's  chisel  had  at  the  top  a  four-armed  projection 
which  played  freely  in  corresponding  slots  of  a  cylinder  attached 
at  one  end  of  the  rope.  As  the  rope  was  lowered  with  its  drill 
it  was  turned  slightly  and  the  ledges  of  the  cylinder  slipped 


DRILLING-  AND  BORING-MACHINES.  59<T 

under  the  cross-arms  of  the  chisel.  This  was  then  caught  on  the 
up- stroke  of  the  rope  and  its  cylinder.  When^it  was  raised  to  the 
proper  height  a  slight  jar  or  jerk  released  the  tool,  whose  cross- 
arms  slid  in  the  slots  and  permitted  it  to  drop  a  distance  equal! 
to  the  length  of  the  slots.  The  invention  of  M.  Hind  and1. 
Chandron  consisted  of  a  pair  of  bent  levers  acting  like  pincers- 
which  grasped  the  tool  and  raised  it  to  the  proper  height,  after 
which  it  was  allowed  a  free  fall  to  do  its  work  without  recoil 
upon  the  apparatus. 

The  jar  universally  used  is  illustrated  in  Fig.  276.  It  is  a. 
pair  of  open  links  allowing  of  i  to  3  feet  of  play.  One  link  is- 
attached  to  the  upper  rod  length  and  its  rope,  the  lower  to  the- 
drill-bit.  These  aggregate  1200  pounds  in  weight  for  an  8"  hole 
and  a  total  drilling  string  of  60  feet. 

The  Drill-tools. — In  Fig.  276  is  illustrated  a  full  string  of 
tools.  Beginning  at  the  left  they  are:  the  sinker-bar,  wrench- 
bar,  wrench,  jars,  temper-screw,  drill-bit,  spudding-bit  and  rope- 
socket,  and  auger-stem.  Above  is  the  floor-circle. 

The  temper-screw  is  suspended  from  the  walking-beam  and 
regulates  the  feed  of  the  drill  and  the  rope.  The  swivel  below 
is  turned  a  half-revolution  after  each  blow,  paying  out  the  rope 
and  rotating  the  drill-bit  a  small  amount.  When  the  full  speed 
of  4  feet  has  been  paid  out  the  screw  is  undamped,  while  the 
drill- rope  is  otherwise  supported,  and  its  feed  raised  for  the. 
next  stage  or  drilling. 

The  stem,  drill-rods,  jars,  and  bit  are  about  ij"  in  diam- 
eter, weighing,  for  a  60- foot  }sring,  1200  pounds.  Each  mem- 
ber has  a  screw-pin  at  its  upper  end,  fitting  accurately  into- 
the  box  at  the  lower  end,  leaving  the  outside  of  the  flush.  A. 
pair  of  wrenches  is  employed  for  tightening,  one  below  and 
the  other  above  the  joint.  The  upper  one  is  further  aided  by 
the  wrench-bar  inserted  into  a  hole  in  the  floor-circle  bolted  to* 
the  frame. 

Drilling  the  Hole. — The  hole  is  started  by  the  spudding-bit 
and  stem,  which  are  attached  to  the  rope  passing  over  the  sheave 
and  once  or  twice  around  the  bull-wheel.  The  engine  in  opera- 


boo 


MANUAL  OF  MINING. 


FIG.  276.— A  String  of  DriU-tools. 


DRILLING-  AND  BORING-MACHINES.  6ox 

tion  revolves  the  bull-wheel  continuously.  The  free  end  of  the 
rope,  being  held  by  a  man,  the  bit  and  rope  are  raised  a  slight 
distance,  after  which  the  rope  is  released,  slipping  on  the  drum, 
and  allows  the  tool  to  fall.  This  is  repeated  till  the  surface-soil 
has  been  penetrated.  The  verticality  of  the  hole  is  maintained 
meanwhile  by  an  upright  plank  box  leader.  Formerly  the 
tool  was  manipulated  by  a  spring-pole  with  a  foot-stirrup  for  av 
moderate  depth  and  was  known  as  "kicking  down  a  hole." 

When  bed  rock  has  been  reached  the  spudding- bit  is  replaced 
by  the  drill-bit,  stem  and  jars,  and  their  rope  clamped  to  the 
temper-screw  which  communicates  the  reciprocating  motion  from 
the  walking-beam.  The  tool  is  raised  from  2  to  6  feet  and  allowed 
to  fall  at  a  rate  varying  with  circumstances,  the  average  being 
six  times  a  minute.  After  each  blow  the  swivel  is  turned,  for  a 
uniform  rotation  of  the  drill  is  essential  and  a  constant  watch- 
fulness necessary,  otherwise  the  drill  may  work  out  of  line..  This- 
is  particularly  essential  if  the  strata  are  inclined  or  the  material 
not  homogeneous. 

Frequently,  to  maintain  verticality  and  sometimes  to  straighten 
a  curved  hole,  a  winged  tool  is  brought  into  play.  The  drill-rod 
is  provided  with  projections  extending  on  the  four  sides  of  the 
chisel  for  the  same  distance  along  it  and  nearly  fitting  the  hole, 
and  sometimes  a  hollow  cylindrical  reamer  is  resorted  to  for  the 
purpose. 

In  the  Mather  and  Platt  system  the  partial  revolution  of  the 
tool  and  its 'flat  rope  was  obtained  after  each  blow  by  a  movable 
collar  rotating  with  the  tool.  It  was  cut  with  inclined  teeth  at 
both  ends  and  was  capable  of  a  motion  of  2"  or  3"  vertically  inside 
of  an  iron  bow.  Above  and  below  the  collar  were  sets  of  inclined 
teeth  into  which  the  collar  ends  geared.  With  each  drop  the 
collar  engaged  the  teeth  of  the  lower  set,  turning  it  with  its  chisel 
one  half  a  tooth.  At  the  end  of  the  upper  stroke  the  collar  strikes 
the  upper  set  and  is  again  turned  one-half  of  a  tooth. 

When  the  temper-screw  has  reached  its  limit  it  is  returned  to 
the  starting-point,  fresh  tools  meanwhile  being  attached;  or  the 
debris  is  pumped  out  preparatory  to  the  next  run.  If  no  under- 


<602  MANUAL   OF  MINING. 

ground  stream  has  been  encountered  to  supply  water  for  cooling 
ithe  bits  and  facilitating  the  removal  of  the  debris,  it  must  be  fur- 
.nished  at  the  surface. 

The  Sand-pump. — This  consists  of  a  slender  cylinder1  with  a 
•valve  at  its  foot  which  is  attached  to  a  rope  having  its  own  pulley 
at  the  derrick  head,  and  its  own  wheel  at  the  engine.  It  is  lowered 
into  the  hole  after  each  run  and  pumped  up  and  down.  The 
sludge  produced  by  the  drill  and  the  water  is  drawn  into  the 
pump  and  raised  to  the  surface. 

Recovering  Lost  Tools. — If  the  sand-pump  or  reamer  breaks 
loose,  it  can  be  fished  out  by  some  form  of  grapnel,  or  it  is  chopped 
up.  If  pebbles  fall  into  the  hole  above  the  tool  and  wedge  it 
fast,  it  may  be  jarred  loose  or  freed  by  a  spear.  Occasionally 
the  hole  is  reamed  out  to  a  larger  size. 

The  Progress. — The  rate  of  progress  varies  from  10  feet  a 
day  in  magncsian  limestone  to  70  feet  in  soft  ground.  Much 
depends  on  the  skill  of  the  driller,  who  may  determine  by  the 
"feel"  of  the  concussion  transmitted  through  the  rope  whether 
the  blow  has  been  effected  or  is  cushioned.  The  feed  is  regu- 
lated accordingly.  Two  men  are  sufficient  to  operate  a  drill 
rig,  and  200  pounds  fuel  are  consumed  per  day. 

The  cost  of  drilling  wells  is  50  and  60  cents  per  foot,  though 
some  very  deep  holes  average  $1.14.  Reaming  is  done  at  a  cost 
of  about  40  cents;  and  the  royalty  for  the  rent  of  a  machine  is 
nearly  10  cents. 

Hand-boring  of  Deep  Holes. — A  unique  mode  of  drilling 
artesian  wells  is  in  vogue  in  San  Luis  Valley,  Col.,  where  300  feet 
of  5"  hole  can  be  made  in  twenty-four  hours.  The  soil  of  that 
.valley  is  very  porous,  and  for  900  feet  down  will  hardly  hold 
water  except  in  the  clay  seams.  Ditches  have  failed  of  their 
irrigating  purposes,  and  each  ranch  is  provided  with  one  or  more 
.of  these  spouting  wells.  From  a  tripod  a  5"  tube  is  held  ver- 
tically and  by  a  pair  of  blacksmith's  tongs  is  turned  by  hand. 
At  the  foot  of  the  pipe  a  bit  projects  outward  £  inch  or  so.  By 
pressure  and  rotation  a  hole  is  cut  spirally  into  the  gravel,  lengths 
i>eing  added  with  the  progress.  A  barrel  or  two  of  water  is 


DRILLING-  AND  BORING-MACHINES.  603 

poured  down  the  pipe  to  wash  the  detritus  out  through  the 
annular  space  until  a  ground-current  is  encountered  which  will 
supply  the  necessary  water.  Beyond  300  feet  horse-  or  engine- 
power  will  be  required  to  turn  the  borer. 

Casing  a  Bore-hole. — When  a  loose  or  fragile  stratum  is 
encountered  its  caving  interferes  with  the  progress  of  the  drill. 
A  pipe-tubing  is  forced  down  to  support  the  sides  of  the  bore. 
It  may  be  driven  while  the  drilling  is  in  progress  below  or  the 
operations  of  drilling  and  casing  may  alternate.  Though  wood 
is  the  best  tubing  material,  the  objections  to  it  are  so  apparent 
that  wrought-iron  pipe  is  used.  The  pipes  may  be  screwed 
together,  joined  by  wrought-iron  couplings  or  telescoped.  The 
first  length,  having  been  sharpened,  is  driven  by  percussion  until 
its  full  length  is  in  the  guide  tube  and  the  successive  lengths 
are  screwed  on  and  hammered.  When  the  depth  has  so  far 
increased  that  the  pipe  can  no  longer  be  rammed,  the  bore-hole 
is  enlarged  for  the  same  size  of  pipe  to  be  lowered,  or  the  same 
diameter  of  bore  is  maintained  for  a  smaller  pipe.  The  pipe  is 
driven  by  repeated  blows  from  a  drop-weight  or  by  the  steady 
pressure  from  jack-screws,  a  block  of  wood  at  the  top  of  the  pipe- 
line receiving  the  impact. 

A  deep  bore-hole  usually  encounters  water-currents  at  differ- 
ent depths  before  the  object  for  which  the  bore  has  been  drilled 
is  attained.  These  are  isolated  by  casing.  A  salt-water  current 
may  be  met  before  the  artesian  flow  is  reached  or  either  of  these 
may  be  intercepted  before  the  oil-sands  are  struck.  A  pipe  of 
large  diameter  may  be  carried  to  the  first  flow  and  sealed  at  its 
foot.  A  second  smaller  pipe,  lowered  to  the  next  flow  and  sealed, 
would  afford  an  escape  for  the  upper  current  in  the  annular 
space  and  a  discharge  for  the  lower  one  inside  the  -smaller  pipe. 
Three  separate  artesian  flows  are  thus  obtained  by  casing  each 
current  and  discharging  them  at  different  surface  elevations.  An 
undesirable  current  may  be  dammed  back  in  like  manner.  This 
may  be  necessary  when  drilling  for  oil,  as  the  accumulation  of 
the  flowing  water  reduces  the  efficacy  of  the  blow  and  also  exerts 
such  pressure,  at  the  rate  of  14.7  Ibs.  per  square  inch  for  every 


604  MANUAL  OF  MINING. 

34  feet  in  depth,  that  when  the  oil  stratum  is  reached  neither 
oil  nor  gas  can  escape. 

The  casing  is  sealed  by  hydraulic  cement  or  by  the  use  of  a 
seed-bag.  A  stout  leather  tube  several  feet  long  is  attached 
outside  of  the  piping  at  the  lower  end.  The  annular  space  is 
filled  with  flaxseed  bran  or  other  absorbent  material  and  the 
top  lashed.  When  lowered  into  place  the  seeds  swell  and  fur- 
nish a  tight  joint  somewhat  like  that  of  the  moss-box  (Fig.  203). 


FIG.  277. — The  Cyclone  Prospecting-machine. 

Shooting  the  Well. — It  frequently  occurs  that  the  oil  does 
not  flow  freely,  or  even  at  all,  when  the  sands  are  encountered.  A 
practice  prevails  of  opening  communication  with  an  underground 
oil-chamber  by  exploding  a  torpedo  at  the  bottom  of  the  well. 
This  is  a  charge  of  nitroglycerine  carefully  lowered  into  position, 
after  which  a  heavy  weight  is  dropped  upon  if.  This  heroic 
treatment  is  not  resorted  to  in  shaly  ground,  unless  provision 
is  made  for  immediately  casing  the  shattered  portion  of  the 
hole,  as  it  often  results  disastrously.  The  charge  varies  from 
10  Ibs.  to  100  Ibs.  in  desperate  cases. 

Removing  the  Casing. — When  the  object  for  which  the  hole 
was  drilled  or  lined  has  been  attained  the  tubing  may  be  recovered 
by  the  use  of  an  ovoid  screw-plug  of  oak,  which  is  attached 
to  the  tool  rod  and  lowered  to  the  bottom.  A  little  sand  is 
poured  in  and  wedges  the  plug,  which  can  be  hoisted  with- 
out trouble.  To  loosen  its  hold  on  the  tube  it  is  only  neces- 
sary to  drop  it  slightly  and  let  the  sand  run  out.  A  three- 


DRILLING-  AND  BORING-MACHINES. 


005 


FIG 


FIG.  279. 

FIG.  278. — The  Keystone  Sand-pump. 
FIG.  279.— Casing  a  Well  in  the  Keystone  Process. 
FIG.  280. — Keystone  Pipe-puller. 


FIG.  280. 


606  MANUAL  OF  MINING. 

pronged  expansible  hook  lowered  down  under  the  tubing  is  also 
called  into  service. 

The  Keystone  Prospecting-machine,  shown  in  Fig.  279,  is 
a  type  of  percussion  drill,  giving  service  in  loose  placer-ground 
and  the  zinc-lead  fields  of  Missouri.  The  string  of  tools  is  iden- 
tical in  character  with  those  of  the  oil-well  rig.  The  hole  is 
started  with  a  fluted  spudding-bit,  stem,  and  rope-socket  and 
alternated  with  the  sand-pump  (Fig.  278).  When  50  feet  has 
been  drilled  the  jars  are  inserted  in  the  line  and 
the  hole  continued. 

The  drive-pipe  casing  is  of  merchant  pipe, 
with  straight  threads  and  squared  ends,  in 
lengths  of  6  feet,  coupled  together.  At  the  foot 
is  a  heavy  drive-shoe  with  a  cutting-edge  and  at 
the  top  a  cap  on  which  the  driving  blows  are 
delivered.  The  casing  follows  the  drilling  closely 
as  shown  in  Fig.  279. 

The  drive-pipe  is  withdrawn  by  the  pipe- 
pullers  (Figs.  280  and  281). 
The  Cyclone  Drilling-machine  resembles  the  other  portable 
drill-rigs,  though  the  detail  of  operation  is  somewhat  different. 
The  drill-bit  is  not  solid  but  is  hollow-stemmed.  Holes  in  the 
side  of  the  bit  afford  communication  from  inside  to  outside. 
Hollow  rods  support  the  drill-bit  instead  of  a  rope  and  at  their 
junction  is  a  ball  valve.  When  the  line  of  tools  falls,  the  valve 
opens  and  permits  the  drillings  to  enter  the  drill-stem  and  rise 
in  the  hollow  rods.  On  the  up  stroke  it  closes.  With  each  down 
stroke  additional  material  enters,  to  be  finally  discharged  at  the 
surface  into  a  proper  receptacle.  Extra-heavy  rod  lengths  are 
employed  near  the  bottom  and  lighter  rods  above.  For  holes 
exceeding  4  inches  in  diameter  the  rods  are  of  2-inch  and  i^- 
inch  pipe  (Fig.  277). 

Rotary  Drills. — M.  Leschot  has  the  credit  for  the  first  appli- 
cation of  rotary  diamond-drills  to  the  miner's  art,  since  which 
time  it  has  gained  in  favor  and  increased  in  range  of  utility. 
The  lower  end  of  a  long  line  of  tubing  is  supplied  with  a  cutting- 


DRILLING-  AND  BORING-MACHINES.  607 

surface  and  is  caused  to  rotate  by  appropriate  machinery  at  the 
surface.  A  continuous  pressure  is  produced  at  the  cutting-edge 
by  the  weight  of  tool  above.  The  cutting-medium  may  be  chilled- 
steel  fragments,  diamonds  inserted  into  the  rim  of  the  boring- 
tool  or  a  steel  tube  may  be  dressed  to  a  hard  cutting-edge  for 
the  purpose,  each  having  its  special  adaptability  to  rock  of  a 
definite  character.  Kl 

The  Diamond-bit. — A  cutting-tool  is  constructed  of  diamonds 
which  are  forced  by  hydraulic  pressure  into  sockets  on  the  end  of 
a  steel  tube.  Recesses  are  accurately  prepared  into  which  they 
are  secured  by  metal  hammered  up  around  them.  In  some 
cases  a  firm  setting  is  obtained  by  forcing  the  stones  forward 
through  small  holes  in  the  metal  by  means  of  a  screw  or  by  hydraulic 
pressure.  A  later  method  consists  in  forcing  the  stones  nearly 
through  the  metal  and  subsequently  grinding  the  steel  down 
until  the  stones  are  exposed. 

The  tube  may  be  closed  by  a  concave  surface  (Fig.  284)  or 
by  a  convex  face  (Fig.  283)  and  the  entire  area  studded  in  such 
manner  that  no  concentric  circle  can  be  drawn  failing  to  touch 
one  or  more  of  the  diamonds.  Some  also  project  beyond  the 


FIG.  282.  FIG.  283.  FIG.  284. 


The  Annular,  Convex,  and  Concave  Diamond-drill  Bits. 

tube.  The  annular  end  of  a  tube  may  be  charged  with  diamonds, 
as  in  Fig.  284.  For  drilling  a  small  hole  the  first-named  types  of 
bit  are  used,  the  concave  being  preferred  to  the  convex,  but 
for  large  holes  or  where  a  central  core  is  sought  and  withdrawn 
for  inspection  the  annular  bit  is  used. 

The  diamonds  are  of  two  varieties,  the  black  "carbonadors" 
and  the  deep-red  "borts."    The  former  lack  the  crystallization 


608 


MANUAL  OF  MINING. 


making  them  durable,  while  the  borts  are  rather  brittle.  Theo- 
retically too  many  carbons  cannot  be  put  in;  "there  should 
be  never  less  than  12,  and  as  many  as  20  may  be  mounted 
on  a  bit." 

The  Diamond-drill.  —  The  bit  is  coupled  to  a  line  of  tubing 
and  guided  by  a  short  or  long  length  of  tube 
(Figs.  285  and  286).  The  tubes  are  in  8-foot 
lengths  and  of  an  outer  diameter  of  rarely  over 
8  inches.  Holes  of  great  depths  are  thus  drilled 
of  a  uniform  diameter  to  the  bottom  or  in  sections 
of  diminishing  diameters. 

The  engine  power  varies  with  the  length  and 
size  of  drill-tube  to  be  manipulated.     For  1000 
*eet  °f  k°re  an  8-horse-power  engine  will  suffice. 
This  with  its  running  gear  is  framed  and  suitably 
braced  to  carry  the  full  line  of  rods.     A  derrick  is  often  erected 


FlG.  286. — The  Core-barrel. 

to  facilitate  the  addition  or  disjointing  of  tubes. 

Regulating  the  Feed  of  the  Drill. — The  changes  in  the  struc- 
ture and  toughness  of  the  rock  during  drilling  are  so  rapid  and 
pronounced  that  the  drill  cannot  be  fed  uniformly.  A  variable 
feed  is  necessary  commensurate  with  the  progress  of  the  drill 
and  two  methods  of  avoiding  a  positive  feed  are  in  vogue.  One 
is  a  spur-wheel  feed,  the  other  an  hydraulic.  The  former  is 
so  adjusted  by  differential  gear  that  its  friction  shall  equal  a 
desired  resistance,  and  when  this  is  exceeded  because  of  undue 
strain  below,  a  regulation  is  obtained.  This  is  inferior  to  the 
hydraulic  feed.  In  Fig.  288  is  shown  a  simple  motor  which  by 
means  of  hydraulic  pressure  on  the  piston  produces  a  pressure 
which  is  maintained  constant.  Both  ends  of  the  cylinder  are 
connected  with  the  pump  and  the  suitable  cocks  admit  of  a  per- 


DRILLING-  AND  BORING-MACHINES. 


609 


feet  control  by  the  operator,  v/ho  gives  any  variation  or  reversal 
of  speed  within  the  limit  of  the  pump  and  the  piston  area. 
When  the  pump  pressure  increases,  the  piston  B  and  its  rods 


FIG.  287. — A  Prospecting  Diamond -drill 


are  raised.  A  decrease  of  "pressure  results  in  its  lowering. 
The  rate  of  feed  is  fixed  by  the  hardness  of  the  rock.  In  Fig. 
287  the  feed  cylinder  is  shown  in  position  above  the  drill-tubes. 
The  pressure  exerted  by  the  feed  is  just  sufficient  to  produce 
abrasion,  but  not  to  cut  the  rock. 


6io 


MANUAL  OF  MINING. 


FIG.  288.— The  Hydraulic  Feed. 


A  chuck  intervenes  between  the 
tube  and  the  feed,  which  may  be 
loosened  at  the  end  of  the  feed-travel 
and  run  up  to  the  top  for  a  fresh 
start.  The  tube  line  is  partially  sus- 
pended by  friction-rollers  at  the  sur- 
face, so  that  it  is  subjected  to  very 
little  tension.  But  a  high  degree  of 
torsion  is  developed  by  the  rotary 
effort  against  the  abrasive  resistance, 
and  the  torsional  strength  of  the  tubes 
places  a  limit  on  the  possible  depth  of 
drilling.  The  regulating  power  of  the 
feed  limits  the  capacity  of  the  ma- 
chine. 

Removing  the  Cuttings  and  the 
Core. — The  tube  is  revolved  at  a  rate 
of  from  400  to  800  turns  per  minute 
and  continues  without  interruption 
from  five  to  eight  hours,  cutting  mean- 
while from  13"  to  2'  each  hour.  A 
stream  of  water  is  fed  inside  of  the 
pipes  and  washes  the  rock  face  and 
the  annular  bit.  It  carries  the  cuttings 
up  through  the  annular  space  outside. 
The  grooves  on  spiral  the  outside  of 
the  pipe  (Fig.  286)  facilitate  the  escape 
of  the  water  with  the  solid  bits.  The 
water  is  fed  outside  the  pipes  and 
escapes  with  the  debris  through  holes 
into  the  interior.  A  guide-bit  (Figs. 
285  and  286)  and  core-barrel  main- 
tain the  bit  in  the  direction  in 
which  it  is  started.  The  water-supply 
is  forced  by  means  of  a  pump. 

The  progress  of  the  drill  is  carefully 


DRILLING-  AND  BORING-MACHINES.  6ll 

watched  and  the  character  of  the  cuttings  supplied  by  the  over- 
flowing water-current  is  examined.  Samples  are  taken  and  pre- 
served for  reference.  With  a  solid  bit  is  produced  only  a  fine 
sludge  for  inspection,  but  the  annular  bit  furnishes  in  addition, 
a  core. 

With  every  10  to   15   feet  of  advance  the  tubes  are   raised 
for  examination  and  for  the   purpose    of   extract- 
ing the  core.     In  the    inside   of   the  drill-tube   is 
a    lifter    (Fig.  289).     This  is  a  spring  ring  which 
grasps  the  core  only  when  it  is  raised     It  wrenches  FIG.  28g.—  The 
the    latter  from    its    place,   and    will    retain    hold     Core-lifter. 
on   it    until    released.      This   can    be    done   by    lowering    the 
tube. 

The  Core. — The  core  furnishes  a  guide  also  as  to  the  direction 
of  inclination  of  the  strata  if  care  be  taken  that  the  tubes  are  not 
turned  while  being  lifted.  Flat  ropes  are  therefore  used  for 
hoisting  as  less  likely  to  untwist  and  turn  the  drills.  False  informa- 
tion may  be  given  by  the  core  if  a  soft  layer  has  been  penetrated 
between  two  hard  strata,  for  the  former  may  wear  away  or  even 
grind  off  loose  in  the  core.  Hence  the  core  may  not  be  an  absolute 
guide.  Many  causes  combine  to  make  this  examination  unre- 
liable. A  careful  measurement,  an  allowance  for  wear,  and 
frequent  inspections  are  the  only  means  of  checking  results. 
The  progress  in  homogeneous  rocks  is  uniform,  but  a  composite 
rock  with  constituents  of  unequal  hardness  is  difficult  to  pene- 
trate in  perfect  alignment.  In  clay  a  thick  pulp  is  formed  which 
chokes  the  tubes  and  requires  the  full  pressure  of  the  pump  to 
remove  it. 

Deflection  of  the  Drill-hole. — Changes  in  the  inclination  of 
the  strata  or  variation  in  hardness  of  the  material  tend  to  deflect 
the  drill-tubes  and  to  buckle  the  rods  under  torsion.  The  correct- 
ness of  the  alignment  must  be  maintained,  and  this  may  be  tested 
by  a  phial  of  gelatine  lowered  to  a  certain  depth  "and  allowed  to 
harden.  When  raised  to  the  surface,  it  is  inclined  in  the  same 
plane  it  occupied  till  the  gelatine  level  is  horizontal,  whence  the 
dip  of  the  hole  is  made  known.  The  line  etched  by  hydro- 


612 


MANUAL   OF  MINING. 


fluoric  acid  on  glass  is  another  method  of  test.     Neither  of  these 
are  entirely  satisfactory. 

Accidents  are  Rare. — Occasionally  a  diamond  may  fall  out, 
and  if  it  cannot  be  recovered,  is  chopped  up  at  once,  or  the  direc- 


FIG.  290. — An  Underground  Diamond  Drill. 

tion  of  the  water  feed  is  reversed  and  its  pressure  increased  to 
wash  the  stone  up  the  tube.  Hard  nodules  of  rock  that  retard 
a  fair  progress  are  chopped  up  by  a  special  bit.  A  diamond  has 
also  been  recovered  by  inserting  a  lump  of  wax  at  the  bottom  of 
a  tube,  which  is  lowered  into  the  hole. 


DRILLING-  AND  BORING-MACHINES. 


613 


The  limiting  depth  of  vertical  holes  is  determined  by  the 
torsonial  strength  of  the  tubes  and  the  power  of  the  machine. 
That  of  inclined  holes  depends  upon  the  amount  of  friction 
between  the  tube  and  the  rock. 

For  underground  work  a  three  horse-power  electric  motor 
like  the  "Little  Beauty"  (Fig.  290)  is  mounted  on  a  truck,  with 


FIG.  291. — Exploration  by  Diamond-drill  Holes. 

drum-drill  and  a  pump,  and  permits  core-drilling  to  advantage 
in  small  spaces.  In  many  mines  i-inch  cores  in  sections  of  5  to 
20  inches  are  cut  for  80  feet  depth,  and  a  great  deal  of  prospecting 
has  been  done  with  this  compact  machine,  which  often  makes 
2  feet  per  hour  at  a  cost  of  68  cents  to  $1.06  per  foot.  In  Fig. 
291  are  exhibited  the  explorations  conducted  at  the  Silver  Islet 
Mine  by  the  use  of  the  underground  drill.  Doubtless  many 
properties  owe  their  existence  to  the  result  of  diamond-drill 
discoveries,  and  its  use  has  frequently  saved  expense  in  various 
ways.  By  graphically  representing  on  cross-section  paper  to 


614  MANUAL  OF  MINING. 

scale  the  results  of  underground  borings  a  more  intelligent  inter- 
pretation can  be  made.  On  the  charts  are  indicated  in  colors 
the  various  rocks  penetrated  by  the  several  bore-holes.  A  com- 
parison of  position  and  thickness  enables  the  engineer  to  judge 
of  the  prospects.  Still,  the  diamond  drill  is  not  considered 
infallible  in  its  indications  as  to  the  presence  or  absence  of  the 
ore  body  sought. 

The  Chapman  hydraulic  pipe-rotating  drill  and  the  Davis 
calyx  drill  are  types  of  rotary  borers  serviceable  for  loose  ground 
such  as  that  of  the  Beaumont  oil-field  of  Texas.  Both  systems 
have  a  turntable  for  directing  the  drills,  and  a  pump  for  the  water- 
supply.  The  Chapman  drill  depends  upon  "adamantine"  for 
its  abrasive  material,  and  the  Davis  drill  uses  chilled  shot  in  hard 
uniform  rock,  but  the  normal  bit  in  soft  ground.  The  drillings 
pass  up  the  tubes  and  are  caught  in  a  calyx  above  the  Davis 
core-barrel,  or  at  the  surface  in  the  Chapman  system.  A  rate  of 
speed  has  frequently  been  attained  equal  to  900  feet  in  twenty- 
five  hours.  In  California  an  experimental  hole  was  drilled  1500 
feet  in  seven  full  working  days.  Holes  of  1 5  inches  diameter  have 
been  bored  by  these  systems  and  to  a  depth  of  3000  feet. 


CHAPTER  VI. 

MINERS'   TOOLS. 

The  Texture  of  Rocks.  — Rocks  to  be  excavated  or  broken 
include  those  which  are  hard  or  soft,  tough  or  brittle,  and  creviced, 
stratified  or  massive ;  and  the  difficulties  of  their  removal  present 
such  varied  and  delicate  questions  that  it  is  difficult  to  give  a 
systematic  account  of  the  principles  of  breaking  ground  or  the 
appliances  employed.  Materials  are  gauged  by  their  resistance  to 
abrasion  being  called  "hard  "  or  "  soft "  as  they  affect  drilling  opera- 
tions; and  tough  or  brittle  according  to  their  resistance  to  concus- 
sion in  blasting.  These  qualities  determine  the  cost  of  breaking 
ground.  For  example,  quartz  is  hard  wherever  encountered, 
but  not  all  of  it  is  tough.  Hard  minerals  wear  the  cutting  tools 
rapidly,  but  a  brittle  mineral  is  easily  shattered  by  the  blow  of 
a  hammer  or  the  concussion  of  a  drill.  A  tough  rock  possesses 
the  tenacity  which  resists  the  rupturing  agency  of  an  explosive. 
It  will  require  powerful  blasting  agents  for  its  removal.  Porphyry 
is  of  an  average  degree  of  consistency  and  may  be  easily  drilled 
and  readily  broken.  Metamorphic  rocks  are  frequently  hard, 
but  always  tough.  They  drill  easily,  but  do  not  break  freely. 
Trap-rock,  syenite,  and  granite  are  hard  on  the  drills,  but  require 
little  explosive,  the  grains  having  slight  cohesion.  Fire-clay  is 
tough,  though  very  soft.  It  has  an  elasticity  which  prevents 
much  impression  being  made  on  it  by  an  explosive. 

The  massive  rocks  are  usually  homogeneous  and  present  the 
same  resistance  throughout  their  mass.  Some  stratified  rocks 
and  gravels  are  not,  being  composed  of  hard  grains  or  pebbles 
cemented  by  a  softer  material.  These  are  easily  ruptured  by 
powder,  but  are  difficult  to  drill  into.  Creviced  rocks  are  not. 

615 


616  MANUAL  OF  MINING. 

economically  broken  by  explosives,  but  are  split  by  wedges  and 
slow  rupturing  agents,  utilizing  the  weak  lines  of  cleavage.  Bitu- 
minous coal  affords  examples  of  a  rock  split  up  by  numerous 
planes  into  prismatic  fragments  which  would  permit  the  gases 
from  powder  explosions  to  escape.  Anthracite  being  brittle, 
of  short  texture  and  hard,  compared  with  bituminous  coal,  there- 
fore requires  powder,  while  bituminous  coal  is  mined  by  picks. 

The  engineer  must  therefore  be  familiar  with  the  texture 
and  behavior  of  the  rock  before  any  estimate  of  the  cost  of  its 
extraction  can  be  made.  In  Drinker's  "Tunnelling"  and  Fos- 
ter's Gallon's  "Lectures  on  Mining"  are  given  elaborate  formulae 
ior  estimating  the  cost  of  extraction,  and  they  may  serve  as  approxi- 
mate guides.  In  each  camp  a  close  observation  of  the  results 
of  experiments  will  supply  the  practical  coefficients  which  can 
foe  used  in  the  above-mentioned  formulae  for  determining  the 
cost  of  working. 

Fire-setting  System. — This  is  a  method  applied  by  the  miners 
of  the  middle  ages  to  conquer  hard  rock  by  the  instrumentality 
of  fire.  The  face  of  the  rock  was  exposed  to  the  heat  of  fire  and 
then  suddenly  cooled  by  water.  The  unequal  contraction,  upon 
cooling,  softened  the  rock,  which  then  was  amenable  to  the  ordi- 
nary tools.  A  grate  was  built  inclined  toward  the  face  of  the 
rock,  and  in  it  were  piled  billets  of  wood.  When  ignited,  the 
flame  was  directed  against  the  breast  by  a  shield  overhead.  In 
some  cases  a  basket  was  suspended  from  the  roof  with  the  same 
Tccult.  Usually  in  the  mines  of  the  early  times  the  wood  was 
fired  on  Saturday  and  allowed  to  burn  out.  The  men  could 
attack  the  calcined  surface  upon  their  return  to  work.  This 
method  may  profitably  be  applied  upon  hard  rock  where  fuel 
Is  cheap  and  the  ventilation  is  good. 

Miners'  Tools. — The  operations  to  be  performed  by  the  miner 
consist  in  breaking  down  the  mineral  or  rock,  removing  the 
1:roken  material,  and  illuminating  the  place  of  work.  These 
require  groups  of  tools  the  nature  of  which  varies  with  the  char- 
acter of  the  mine  in  which  the  work  is  performed.  The  mineral 
may  be  broken  down  by  picks,  drills,  striking-hammers,  sledges, 


MINERS'   TOOLS.  617 

augers,  gads,  moils,  wedges,  and  feathers.  Auxiliary  tools  are 
also  employed,  such  as  the  swab,  spoon,  gun,  etc.,  when  the  mate- 
rial is  tough  enough  to  require  blasting.  In  the  latter  case,  in 
addition  to  the  striking- hammers  and  drills,  will  also  be  required 
the  tools  for  making  powder  cartridges,  cans  for  carrying  the 
same,  and  needles,  squibs,  fuses,  etc.,  to  ignite  the  powder.  The 
mineral  is  removed  by  the  aid  of  shovels,  picks,  and  spades; 
occasionally  sledges  and  crowbars  are  required  for  loosening  or 
breaking  fragments  too  large  for  convenient  handling.  The 
illumination  of  the  place  of  work  will  require  a  candle  and  its 
stick,  or  oil-lamp  and  an  oil  supply. 

All  tools  loaned  or  handed  to  the  miners  are  entered  on  an 
account  against  them,  as  otherwise  many  would  be  lost  in  the  waste, 
or  otherwise  disappear.  A  check  is  made  of  the  condition  of  all 
tools  returned  by  the  men.  Steel  should  be  measured  or  weighed 
out  on  delivery  at  the  beginning  of  the  day,  month,  or  contract. 

Rotary  Drills. — The  second  class  of  drills  cut  away  the  rock 
by  the  rotary  motion,  and  according  to  the  character  of  the  cut- 
ting bit  they  are  breast-augers,  roof-augers,  and  machine-power 
augers,  worked  by  compressed  air  or  electricity. 

Hand  Boring-machines  consist  of  a  drill-piece  of  twisted  steelr 
W7elded  to  an  iron  stem  and  fixed  to  the  end  of  a  long  screw  pass- 
ing through  a  nut  which  is  supported  on  an  upright  bar  by  a  ball- 
and-socket  or  universal  joint,  permitting  any  direction  of  bore. 
The  rotary  motion  is  ccmmunicated  to  the  drill-screw  by  two 
bevel-gear  wheels,  or  directly  by  a  revolving  handle  at  the  rear 
of  the  screw.  The  bit  may  be  a  wedge-shaped  bit  with  a  V  edge, 
or  it  may  be  twisted  into  the  common  auger  form. 

The  form  illustrated  in  Fig.  292  is  a  special  bit  used  by  bitu- 
minous miners.  The  shank  to  which  the  twisted  steel  bit  is 
welded  is  bent  to  form  a  handle  and  terminates  in  a  breastplate 
(Fig.  293).  The  boring  of  the  holes  is  accomplished  by  hand, 
the  pressure  being  applied  by  the  weight  of  the  body  against  the 
plate.  The  bits  of  these  augers  are  from  ij  to  i\  inches  in  diam- 
eter. Sectional  augers  are  more  convenient  for  delivery  to  the 
blacksmith,  therefore  they  are  made  removable  from  the  shank. 


6i8 


MANUAL  OF  MINING. 


Of  the  hand-  and  roof-power  augers  there  are  several  styles, 
the  two  general  classes  being  the  grip-drill  and  the  post-drill. 
The  grip-drill  (Fig.  294)  has  a  metal  rest  a  for  the  feed-screw 
nut  &,  in  which  the  feed-screw  c  turns  while  the  hole  is  being 
bored.  A  small  recess  in  the  coal  is  provided  into  which  the 
grip-bar  d  is  driven  and  wedged  and  thus  gives  stability  to  the 
machine.  The  boring  can  be  done  from  the  side  or  end. 


FIG.  292. — A  Bituminous-coal  Bit. 


osoooc=i/          \^ 


L 


FIG.  294. — A  Grip-drill. 

The  post- drill  consists  of  a  suitable  borer  attached  to  the 
post,  which  is  usually  placed  upright,  having  a  jack-screw  at  one 
•end  and  firmly  wedged  between  the  floor  and  the  roof.  The  post 
is  split  or  may  be  obtained  solid.  The  drill  is  attached  to  the 
feed-screw  which  advances  through  the  feed  end  by  a  handle 
and  crank  operating  the  gearing  shown.  The  universal  joint  sup- 
porting the  drill-gear  permits  holes  to  be  drilled  in  any  direction. 


MINERS'    TOOLS. 


619 


Portable   Power   Borers. — The   Jeffrey   rotary   electric    coal- 
drill   (Fig.  295)  has  a  small  enclosed  motor  on  the  end  of  whose 


FlG.  295.— The  Jeffrey  Rotary  Electric  Coal-drilL 

armature-shaft  is  a  small  pinion-gearing  with  a  larger  wheel. 
Two  drills  are  used  for  each  hole,  one  being  3  feet  and  the  other 
6  feet.  The  time  required  to  drill  a  6-foot  hole  in  ordinary 
bituminous  coal,  with  either  type,  is  from  twelve  to  fourteen 
minutes.  In  addition  to  gearing  for  hand-power  drills  there 
are  ratchet  attachments  for  driving  them  in 
such  places  where  the  holes  must  be  drilled  too 
close  to  reach  the  roof  or  the  floor  for  the  crank- 
motor  to  be  used.  It  requires  from  |  to  i  horse- 
power to  drive  them.  The  power  is  communi- 
cated through  gears  and  a  small  pinion  by  which 
rotation  is  produced.  A  feather  on  the  larger 
gear-wheel  slides  along  a  longitudinal  slot  in  the 
feed-screw  and  permits  of  an  advance  of  the  bit 
with  the  progress  of  the  boring.  These  force  the 
drill  forward  at  any  determined  rate  without  un- 
duly crowding  the  bit. 

A  similar  pattern  of  drill  (Fig.  296)  is 
operated  by  a  small  air-engine  or  by  hand.  In 
a  hand-borer  this  drill  is  serviceable  where  electric  or  air  coal- 


620  MANUAL  OF  MINING. 

• 

cutters  are  also  employed.  Unless,  however,  the  power-plants  are 
to  be  installed  for  other  purposes,  it  is  hardly  necessary  to  install 
one  safely  for  the  use  of  these  drills,  so  long  as  equally  serviceable 
hand-power  drills  can  be  obtained. 

In  loose  material  and  shattered  rock,  shovels  and  spades  are 
the  only  tools  needed;  clays  and  soft  rocks  require  picks,  crow- 
bars, and  shovels;  ground  that  is  scaly,  brittle,  and  seamy  is 
split  by  wedges  in  different  ways.  Massive  rock  cannot  be 
broken  without  the  hammer  and  drill.  Powder  and  explosives 
arc  used  in  the  latter  case,  and  occasionally  in  the  others,  as 
auxiliary  aid.  For  economy  and  speed,  there  are  machines  which 
are  driven  by  steam,  electric,  or  pneumatic  power,  imitating  the 
operations  of  the  above  hand  tools. 

Underholing  Bituminous  Coal. — The  operation  called  under- 
cutting, underholing,  bearing  in,  or  kirving,  consists  in  cutting  a 
groove  underneath  a  mass  of  coal  in  the  soft  floor,  or  in  the  coal 
itself,  to  a  depth  about  equal  to  the  thickness  of  the  bed.  The 
groove  is  9  inches  high  at  the  face  and  about  2  inches  at  the  back. 
The  coal  is  propped  by  sprags  while  the  miner  is  engaged  in 
digging  a  deep  groove.  The  coal  then  freed  below  may  be 
broken  off  by  the  pressure  of  the  roof  overnight,  or  its  fall  may  be 
furthered  by  a  few  drill-holes  loaded  with  powder.  The  block 
may  also  be  sheared  by  grooves  along  the  sides  of  the  rooms  or 
stalls,  cut  vertically  from  roof  to  floor.  Sometimes  all  three 
methods  are  practised  to  break  the  coal  down.  Shearing  is 
more  expensive  than  undercutting,  hence  is  used  only  for  shallow, 
firm  coal  or  under  a  very  strong  roof.  The  cleats  in  the  coal 
materially  assist  in  its  breaking.  Anthracite  is  too  firm  to  be 
treated  in  this  way;  it  must  be  blasted  off  the  solid. 

Shooting  Off  the  Solid. — This  term  is  employed  to  designate  the 
character  of  blasting,  which  is  in  practice  in  the  anthracite-coal 
mines  wherever  there  is  but  one  face  exposed  toward  which  a 
blasting  agent  may  operate.  In  anthracite  mines  this  practice 
is  universal,  the  miner  exploding  the  holes  so  that  each  succeeding 
one  will  have  a  free  face  for  blasting.  In  flat  beds  he  makes 
centre  cuts  (Chapter  VII)  on  the  face  by  blasting  wedged-shaped 


MINERS'    TOOLS.  621 

masses  of  coal.  In  pitching-beds  he  blasts  in  the  coal  near  the 
floor  and  undermines  the  bed.  By  systematic  work  a  keg  of 
25  Ibs.  of  powder  can  thus  blow  down  from  30  to  50  cars  of  coal. 
This  practice  is  warranted  in  bituminous  mines  when  the  coal 
is  to  be  delivered  to  the  coke-ovens,  but  not  when  it  is  tc/  be 
marketed  for  domestic  use. 

Shovels. — The  ordinary  shovels  are  made  of  iron  plate  rolled 
under  a  welding  heat  with  an  edge  of  steel,  and  ears  drawn  out 
for  the  handle.  A  concavity  to  the  blade  imparts  stiffness  and 
carrying  capacity  to  the  tool.  Such  a  scoop-shaped  shovel  is 
preferred  by  the  anthracite  miner.  The  end  of  the  shovel  is 
square  or  pointed,  according  to  its  uses.  The  pointed  blade 
with  the  long  handle  is  the  most  used.  That  of  the  "diamond 
label"  is  the  most  popular;  the  width  of  blade  in  inches  desig- 
nates the  commercial  number  or  size  of  the  shovel. 

Picks. — The  pick  is  variously  known  as  a  pike,  mandril, 
slitter,  hack,  and  mattock,  according  to  its  shape  and  length. 
It  is  of  iron  with  steel  tips  or  all  steel.  It  has  an  opening  or  an 
eye  in  the  centre,  the  ends  being  pointed  to  a  pyramidal  shape 
or  a  chiselled  edge.  The  points  or  tips  are  hardened  to  suit  the 
nature  of  the  work  to  be  done.  They  are  the  wearing  parts  and 
should  be  replacable.  On  this  account  the  all-steel  pick  is  not 
serviceable.  Each  sharpening  after  use  consumes  an  amount  of 
metal,  which,  in  time,  reduces  it  to  an  inconvenient  length.  An 
iron  pick  14  inches  long  with  two  steel  tips  wedged  into  the  ends 
gives  longer  service,  for  the  points  when  worn  out  may  be  replaced. 
Removable  points  are  sometimes  supplied  with  picks.  "Pick 
steel"  is  a  special  steel  in  bars  i£"Xf"  or  i£"Xf"  which  are 
cut  into  suitable  lengths. 

The  picks  are  generally  made  at  the  mine  by  welding  into  the 
iron  lengths  of  steel  wedged  and  pointed.  When  finished  they 
are  about  22  inches  long,  though  some  are  29  inches  long,  and 
in  the  hands  of  the  "box"  cutters  are  doing  remarkable  "jadding" 
(cutting  the  top).  The  weight  of  picks  varies  between  2  and 
9  Ibs.,  with  3 1  or  4  Ibs.  as  an  average,  the  heavier  weights  being 
used  for  downward  cutting. 


622  MANUAL  OF  MINING. 

The  eye  of  the  pick  is  oval.  It  is  formed  by  gashing  the  red- 
hot  bar  in  the  middle,  upsetting  it,  cutting  it  open  by  a  drift,  and 
hammering  out  stout  cheeks  with  abundant  metal  at  the  sides 
and  a  bearing  surface  for  the  handle  as  long  as  possible.  All 
of  the  strain  of  prying  by  the  points  and  handle  falls  upon  the  eye, 
which  therefore  must  be  stout. 

The  shape  and  weight  of  the  pick  depends  on  the  service  it 
is  to  give.  There  is  the  straight  or  the  curved  pick,  the  anchor- 
or  the  poll-pick,  each  having  favor  for  certain  work.  Indeed, 
in  the  same  mine  several  forms  are  to  be  seen,  perhaps  as  a  matter 
of  individual  prejudice.  The  straight  pick  assists  the  reach,  the 
curved  pick  enables  a  more  effective  blow  to  be  struck.  So, 
for  overhand  work,  for  underholing,  and  for  getting  into  corners, 
a  straight-head  pick  is  used,  but  does  not  strike  as  well  in  down- 
ward work  as  the  curved.  The  curve  of  the  latter  should  be  an 
arc  of  a  radius  equal  to  the  combined  lengths  of  the  arm  and 
handle. 

Usually  the  bituminous-coal  miner  has  a  light  pick  of  2  or 
3  Ibs.  weight  for  the  clay  and  a  heavier  one  for  cutting  bony 
coal,  slate,  etc.  The  light-weight  picks  have  one  square  chiselled 
edge  and  a  pointed  one.  The  anthracite  picks  are  heavier.  The 
poll-pick  used  for  prying  loose  coal  from  the  face  or  rock  from 
the  roof  is  as  heavy  as  can  be  conveniently  handled. 

The  pick  is  single-  or  double-pointed.  In  the  first  variety  the 
other  end  is  forged  into  a  hammer-head,  known  as  a  "poll,"  and  is 
used  for  breaking  fragments  or  driving  a  wedge.  The  tapered  or 
chiselled  end  of  the  pike  is  used  to  pry  into  and  slit  the  rock  with 
the  leverage  obtained  from  the  length  of  handle.  For  soft  ground 
the  pointed  end  is  long  and  slender. 

The  handle  or  helve  of  hickory  or  ash  is  a  straight-grained, 
firm,  well-trimmed  stick.  It  is  trimmed  to  the  shape  of  the  eye 
and  wedged  tight  by  a  pair  of  iron  feathers.  The  handle  should 
be  at  right  angles  to  the  pick.  When  so  fastened,  the  arcs  drawn 
by  the  tips,  using  the  handle-ends  as  centres,  will  have  equal  radii. 
The  length  of  American  handles  is  about  28  inches;  those  of  the 
English  are  30  to  35  inches. 


MINERS'    TOOLS.  623 

Breaking  Dimension  Stone. — Quarrying  at  the  surface  involves 
the  procural  of  blocks  of  stone  of  definite  shape  and  size  suitable 
for  dressing  by  the  stone-mason  to  desired  dimensions.  The 
process  must  be  such  as  to  ensure  with  some  degree  of  accuracy 
the  shape  desired  without  injury,  and  the  rupturing  agents  must 
be  less  violent  than  powder.  Wedges  were  therefore  used,  though 
they  are  rapidly  becoming  obsolete.  The  primitive  wedges  were 
of  dry  or  wet  wood.  Now  they  are  of  steel  or  iron.  They  were 
forced  or  hammered  into  the  planes  of  cleavage  of  the  rock,  splitting 
the  latter  along  the  given  directions. 

In  Fig.  306  is  illustrated  a  block  being  split  by  a  number  of 
plugs.  The  wedges  are  driven  by  a  sledge  in  succession,  begin- 
ning at  either  end  and  working  toward  the  centre,  each  one  receiv- 
ing a  few  taps  in  turn.  The  holes  are  put  alternately  long  and 
short,  and  of  a  depth  12  to  20  inches,  according  to  the  thickness 
of  the  stone. 

The  rock  may  also  be  broken,  by  forming  a  rift  along  the 
face,  by  a  sledge.  The  stone  is  then  pounded  along  a  given  line 
rapidly,  until  the  continual  concussion  overcomes  the  tenacity  of 
the  stone,  and  starts  a  crevice  along  the  line  into  which  a  wedge 
may  be  inserted  and  driven. 

The  Plug  and  Feather  are  more  efficient  than  the  simple 
wedge.  The  plug  is  a  flattened  wedge  and  the  feathers  have  two 
wedges  each,  with  one  face  flat  and  the  other  curved.  They  are 
introduced  into  a  drilled  hole  with  their  points  up.  The  plug 
is  hammered  between  them  and  the  design  is  such  that  the  feathers 
are  forced  apart  at  the  bottom  of  the  hole  with  greater  power 
than  at  the  top. 

Hydraulic  Wedges. — Hydraulic  wedges  have  been  used  with 
good  success  at  the  collieries  near  Saarbruecken,  the  position  of 
the  driving- wedge  being  reversed  from  that  of  the  plug,  the  thicker 
end  being  placed  in  the  bottom  of  the  hole  and  the  edge  near  the 
top.  Fitting  between  the  two  half-round  wedged  cheeks  which 
point  down,  it  is  driven  from  below  upward  by  the  hydraulic 
pressure,  or  the  force  from  the  explosion  of  powder.  This  snaps 
the  rock  in  the  plane  of  the  thin  edge.  Dimension  store  is 


624 


MANUAL  OF  MIX  IXC. 


obtained  in  some  quarries  by  this  process,  but  the  risk  of  huffing 
off  the  rock  at  the  top  of  the  hole  is  great. 

Gads  or  Moils  are  very  useful  accompaniments  to  every 
miner's  equipment,  differing  from  the  early  wedges  only  in  that 
they  have  pointed  tips,  not  chiselled  edges.  A  piece  of  steel  that 
has  done  service  for  drilling,  has  been  dressed  and  smithed  until 
it  is  less  than  10  inches  long,  and  can  no  longer  be  used  for 
starting  a  hole,  is  converted  into  moil  by  tapering  off  the  bit- 
point.  It  is  used  for  chipping  or  trimming  the  rock,  for  a  smooth 
bearing  to  timbers,  etc.  It  is  used  in  the  Lake  Superior  copper- 
mines  for  "blockholing"  or  splitting  the  large  masses  of  copper. 
It  is  almost  indispensable  in  hard  rock,  an'd  a  number  of  them, 
sharpened  and  hardened,  should  always  be  on  hand  for  the  use  of 
timbermen  and  shaftsmen. 

A  crowbar  is  a  tool  of  occasional  service  underground.  It  is 
a  long,  unhardened,  steel  bar  with  a  point  at  one  end,  and  serves 
as  a  pry. 

Jumpers  and  Drill-rods. — Percussion  on  the  rock  is  accom- 
plished by  a  jumper  for  vertical  holes  only  or  by  a  drill  for  holes 


(b) 

FIG.  297. — Forms  of  Bits. 


of  any  angle.  These  are  heavy  steel  bars  provided  with  cutting 
vH<*es,  which  may  be  straight,  concave,  or  convex,  and  acute 
(slim)  or  "bluff"  (Fig.  297),  the  shape  being  somewhat  a  matter 
of  individual  preference  and  skill.  The  bit  is  shaped  by 


MIXERS'    TOOLS.  625 

the  blacksmith,  who  hammers  the  red-hot  end  of  a  steel  drill- 
rod  to  a  wedge-shape,  flaring  it  out  to  a  width  somewhat 
greater  than  that  of  the  rod.  The  edge  is  worked  to  a  straight 
or  curved  form,  and  then  hardened  by  reheating  the  end  to 
the  proper  degree  and  plunging  it  into  a  cooling  fluid.  The 
convex  bit  is  stronger  than  the  straight  bit,  and  transmits  the 
blow  better  with  less  risk  of  fracture  at  the  corners.  The 
cutting  edge  of  |-inch  drills  is  i|  inches  wide,  and  other  sizes 
are  correspondingly  flared.  The  other  end  of  the  drill-rod  is 
not  dressed  after  being  cut  off.  It  receives  the  blow  of  the 
striking  hammer.  When  both  ends  of  the  drill-rod  are  worked 
with  cutting  edges  the  tool  is  known  as  a  jumper. 

The  Jumper  is  a  heavy  5-  or  6- foot  round  iron  bar,  heavier 
near  the  middle  and  hammered  to  an  edge  at  each  end.  A  hole 
is  churned  down  into  the  rock  by  lifting  the  jumper  and  allowing 
it  to  fall,  after  which  it  is  turned,  slightly  raised,  and  dropped  a 
foot  or  so.  When  one  bit  is  dulled  the  jumper  is  reversed  and 
the  other  end  continues  the  operation.  In  this  way  a  hole  is 
"jumped  down."  The  debris  is  cleaned  out  at  intervals  by  a 
scraper  or  a  spoon.  Coal  and  limestone  are  churned  in  this  way 
at  a  rate  of  40  or  50  feet  of  hole  in  a  shift;  granite,  at  15  feet. 
The  jumper,  with  one  cutter  and  a  head  for  hammering,  has  lost 
the  significance  of  its  name  and  is  properly  a  drill. 

The  Drill. — The  steel  rods  for  hand-  or  power-drills  are  pro- 
cured in  lengths  of  14  feet  and  of  any  diameter.  Though  there 
are  many  grades,  the  English  Jessop  and  the  American  brand 
"Black  Diamond"  have  pronounced  adherents.  The  steel  is 
round  or  octagonal,  the  diameters  usually  being  between  f  and 
2  inches.  The  length  of  the  bar  depends  upon  the  depth 
of  the  hole  to  be  drilled.  The  rods  of  less  than  i  inch  are  used 
only  for  short  holes  and  light  hammering,  as  when  a  miner 
both  holds  the  drill  and  strikes  it  with  a  hammer.  In  "double- 
hand  work,"  where  a  companion  strikes,  the  diameter  of  the 
drill  is  between  i  and  i|  inches.  For  "three-hand"  or  "double- 
hammer"  work  the  steel  may  be  as  large  as  2  inches,  though 
it  is  doubtful  if  a  hand-drilled  hole  greater  than  i|  inches 


626  MANUAL  OF   MINIXG. 

diameter  is  justifiable.  Generally  speaking  it  is  preferable  to 
have  the  drill  as  large  as  possible  to  transmit  the  blow,  and  also 
to  produce  a  hole  as  broad  as  convenient.  The  drill  of  small 
diameter  is  rather  light  and  yields  under  the  blow  without  trans- 
mitting the  effect. 

Drilling  Operations. — The  hole  is  drilled  by  the  repeated 
percussive  blows  upon  the  drill  by  the  hammer.  The  rock  is 
chipped  by  the  concussion  of  the  bit;  after  each  stroke  the  drill 
is  turned  slightly  and  the  blow  repeated.  Though  the  hole 
may  be  drilled  without  the  addition  of  any  water,  it  is  never- 
theless the  practice  to  occasionally  introduce  some  water  into 
the  hole  to  preserve  the  temper  of  the  tool  and  to  facilitate  the 
removal  of  the  debris. 

The  lengths  of  drills  employed  increase  with  the  depth  of 
the  hole  and  its  progress,  the  starting  drill  being  a  short,  stout 
drill  of  comparatively  large  diameter,  and  the  finishing  drill,  30 
to  36  inches  in  diameter,  being  of  smaller  diameter.  The  several 
lengths  of  drills  constitute  a  set,  the  bits  of  a  set  being  flared  less 
with  each  increasing  length  of  drill. 

The  depth  of  holes  varies  with  the  character  of  the  work.  It 
rarely  exceeds  36  inches  with  double-hand  work  and  averages 
25  inches  in  single-hand  work.  In  hard,  brittle  rock  the  holes 
are  made  deep  and  narrow,  while  in  the  tough  and  fissured  mate- 
rial they  are  short  and  broad.  Deep  holes  are  more  economical, 
but  the  miner  usually  refrains  from  drilling  to  the  limit,  except 
upon  rare  occasions,  because  of  the  difficulty  arising  from  leaving 
the  ground  in  poor  shape  for  subsequent  blasting,  as  will  be 
seen  in  the  Chapter  on  Blasting.  The  miner  places  the  direc- 
tion of  the  holes  and  carries  them  to  such  a  depth  as  will  enable 
him  not  only  to  produce  a  maximum  effect,  but  also  to  leave  a 
suitable  free  face  for  attack  by  the  subsequent  shots.  As  to 
diameter,  it  is  cheaper  to  drill  a  small  hole  and  to  increase 
the  strength  of  powder  than  to  have  a  hole  of  large  diameter 
with  a  large  quantity  of  weak  powder.  The  work  of  drilling 
a  ij-inch  hole  is  nearly  three  times  that  of  putting  down 
a  ^-inch  hole  of  the  same  length;  the  relative  volumes  of 


MINERS'   TOOLS.  627 

the  rock  pulverized  are  as  the  square  of  the  diameters;  or, 
in  other  words,  all  else  being  equal,  25  lineal  feet  of  small 
holes  can  be  put  down  while  9  feet  of  the  i|-inch  hole  are 
being  drilled. 

Twelve  feet,  in  the  aggregate,  constitute  a  stent  with  double- 
hammer  work,  and  9  feet  of  holes  with  the  single  hammer  and 
one  striker.  In  the  case  of  single-hand  work  the  bits  are  smaller 
in  diameter  than  in  the  double-hand  work  for  a  gang  of  two 
strikers.  About  30  inches  of  holes  can  be  drilled  per  shift  by 
single-hand  work  in  rock  of  a  medium  character.  Jn  rare  excep- 
tions the  progress  may  be  as  great  as  5  feet  per  man. 

The  consumption  of  steel  varies  with  the  tool,  but  an  average 
may  be  given  as  about  i  Ib.  for  15  tons  of  rock  blasted. 

Removing  the  Drillings. — During  the  progress  of  drilling 
the  debris  is  removed  by  a  squib,  stick,  spoon,  or  gun,  accord- 
ing to  the  circumstances.  The  spoon  is  a  round  f-inch  iron 
bar,  40  inches  long,  having  a  handle  at  one  end  and  a  spoon  at 
the  other,  which  latter  is  made  by  curving  slightly  the  lower  5  or 
6  inches  of  its  length  and  bending  up  a  spoon  at  the  end.  The 
gun  is  a  syringe,  made  of  a  length  of  gas-pipe  and  fitted  with  a 
suction-piston  having  a  handle  on  its  rod.  By  these  appliances 
the  load  may  be  removed  from  the  hole  more  effectually  than 
can  be  done  by  the  spoon.  The  latter  is  employed  for  scraping 
the  sand  from  a  dry  hole.  The  swab-stick  is  sometimes  used 
to  remove  the  last  remnant  of  moisture  in  the  hole  prior  to 
charging  it  with  powder. 

Hammers  are  employed  for  breaking  large  fragments  of 
rock  and  for  striking  drills.  In  weight  and  finish  of  their  faces 
they  differ.  The  face  is  usually  flat  and  very  hard.  The  convex 
striking-face  is  rarely  used.  Striking-hammers  are  not  used  for 
rock-breaking,  which  duty  is  left  to  the  sledges. 

For  single-hand  work  the  hammer-head  weighs  3^  to  5  Ibs.; 
for  double-hand  work,  5  to  7  Ibs.,  except  for  wedge- driving;  it  is 
short  with  a  broad  face,  with  a  handle  correspondingly  short  or  long. 

Single-hand  and  Double-hand  Work. — By  single-hand  work 
is  understood  the  method  of  drilling  holes  by  a  single  individual, 


628  MANUAL  OF  MINING. 

who  holds  the  drill  in  one  hand  and  strikes  it  with  a  hammer  in 
the  other.  In  double-hand  work  the  operations  of  holding  and 
turning  the  drill  and  of  striking  the  tool  are  performed  by  two 
men.  In  double-hammer  work  the  gang  is  composed  of  three 
men — two  striking  and  one  holding.  Except  in  very  hard  rock, 
in  which  satisfactory  progress  is  made  by  double-hammer  work, 
single-hand  work  is  the  usual  practice  in  narrow  places  and 
double-hand  work  where  the  room  is  sufficient  to  allow  of  swing- 
ing the  sledge  for  a  blow.  For  many  reasons  the  double-hand 
work  is  almost  universal.  Though  many  Abjections  obtain  to 
its  practice,  it  has  an  advantage  in  enabling  the  men  to  alter- 
nate the  work  of  striking  and  holding  and  thus  relieve  one  another. 
But  the  benefits  derived  from  the  single-hand  work  are  so  great 
that  where  it  is  possible  of  introduction  it  is  employed. 

Drinker  says  that  in  the  point  of  economy  of  time  and  money 
"one-hand  drilling"  is  from  30  per  cent  in  soft  schist  to  20  per 
cent  in  soft  sandstone  cheaper  than  "two-hand  drilling."  In 
hard  rock  "one-hand  drilling"  gives  the  more  rapid  advance. 
Dr.  A.  Serlow,  in  his  "Leitfaden  zur  Bergbaulunde,"  believes 
that  except  in  shaftw'ork  all  other  forms  of  drilling  may  be  exe- 
cuted more  accurately  by  the  "  single  "  than  by  the  "  double-hand, " 
and  perhaps  more  cheaply. 

Itemizing  the  Drilling  Expenses. — There  are  various  methods 
of  apportioning  the  cost  of  broken  ground.  One  simple  method 
consists  in  keeping  the  supplies,  such  as  drill  steel,  tools,  etc., 
under  one  heading;  the  power,  fuse,  caps,  oil,  and  candles  as 
sundries;  the  labor  by  itself;  while  the  fixed  charges,  such  as  office 
and  shipment  expenditures,  are  given  separate  accounts.  Occa- 
sionally the  labor  item  is  subdivided  and  distributed  as  a  guide 
for  future  economies.  It  may  be  possible  thereby  also  to  ascer- 
tain the  number  of  tons  or  cars  broken  per  miner,  from  which 
may  be  fixed  a  standard  of  settlement.  Again,  the  cost-sheet  may 
be  divided  into  accounts  as  follows:  Under  labor  are  miners, 
drilling,  blasting,  trammers,  blacksmiths,  engineers,  timbermen. 
trackmen,  engineering,  and  superintendence.  The  minerals- 
account  contains  items  for  powder,  caps,  and  fuse,  oils,  waste, 


MINERS'    TOOLS.  629 

candle,  coal,  stable  repairs,  track  tools,  and  timber;  and  the  mis- 
cellaneous account  includes  office  expenses,  taxes,  insurance, 
sinking-fund,  interest,  and  shipments.  These  different  items  are 
reduced  to  cost  per  ton  of  product  as  well  as  the  aggregate  cost 
per  week  or  month. 

The  Blacksmith  Shop. — The  duties  and  work  of  the  blacksmith 
may  not  seem  relevant  to  the  engineer,  but  no  manager  can  afford 
to  be  ignorant  of  any  element  connected  \vith  the  economy  of  his 
work.  As  a  matter  of  fact,  it  is  highly  essential  that  the  latter 
be  capable  of  judging  the  performance  of  the  blacksmith,  who 
is  the  butt  between  the  complaints  of  a  miner's  inefficiency  and 
that  man's  retorts  in  pleading  bad  tools.  A  smith  who  can 
sharpen  tools  suitably  for  hard  ground  is  held  in  high  esteem  by 
miners.  The  equipment  of  the  shop  comprises  a  full  kit  of 
tools,  costing  perhaps  $30.00.  a  good  bellows  and  tuyeres,  Peter's 
anvil,  vise-taps  and  dies,  twist-drills  and  hoop  iron,  an  assort 
ment  of  carriage-  and  machine-bolts,  screws,  spikes,  nails,  a  few 
horseshoeing  tools,  benches,  etc.  The  building  is  about  14'  Xi2f 
with  hinged  door-openings,  near  or  over  the  fire  and  in  the  two 
walls,  for  working  long  bars.  If  machine-drills  are  used  in  the 
mine,  a  set  of  special  swages  are  necessary  for  finishing  the  bits 
to  the  X,  Z,  or  +  form,  as  required. 

The  miners,  who  usually  must  pay  for  the  wasted  steel  or  for 
the  sharpening  of  bits,  are  entitled  to  the  services  of  a  competent 
blacksmith  and  suitable  drill  steel.  The  charge  is  usually  $i 
per  month  in  coal-mines  and  3  cents  per  bit  in  metalliferous 
mines.  The  sharpening  is  often  done,  even  for  contractors, 
at  the  expense  of  the  mine. 

For  a  small  mine  employing  20  men  in  all,  i  blacksmith 
will  suffice,  though,  of  course,  it  depends  upon  what  he  must 
do.  A  good  sharpener  can  dress  tools  for  20  men  on  medium 
rock,  or  swage  the  I  or  X  bits  for  7  machine-drills.  Excepting 
the  pointing  of  picks,  the  cutting  of  steel,  and  the  handling  of 
large  pieces,  he  will  need  no  striker.  With  this  help  he  can 
make  12  heavy  picks,  20  light  ones,  or  weld  40  pick-stems  in  a 
shift ;  or  he  can  finish  2  sets  of  colliers'  tools  of  5  coal-picks} 


630  MANUAL  OF  MINING. 

2  wedges,  a  hammer,  and  2  bottom-picks.  Alone,  he  can  dress 
40  bits  an  hour;  with  help,  he  can  forge  25  double  hand-bits, 
or  draw  out  and  temper  50  pick-points  per  hour. 

Forge  Fuel. — One  important  element  of  success  in  forging 
iron  is  a  clean,  pure  fuel.  This  may  be  a  slightly  caking  coal 
that  gives  flame  and  a  high  heat,  or  coke  which  is  hotter  but 
more  difficult  to  keep  alight.  The  coal  should  be  clear  of  shale 
and  slate,  for  they  fuse  and  make  a  pasty  cinder  that  adheres  to 
the  iron.  It  must  be  free  of  sulphur,  which  makes  the  iron  "hot 
short"  and  also  tends  to  produce  scales  while  forging.  For  the 
latter  reason  white-ash  coal  is  preferred  to  red-ash  coal. 

Welding. — A  very  useful  property  of  wrought  iron  is  its 
capability  of  welding,  by  which  two  short  lengths  may  be  united 
to  form  a  bar  of  a  serviceable  length.  The  process  consists  of 
wedge-tapering  an  end  of  each  bar,  heating  them  to  red,  and  sub- 
sequently hammering  the  softened  parts  together.  A  more 
difficult  joint,  known  as  the  split,  is  described  further  on  in  the 
steeling  of  picks.  If  the  welding  has  been  well  done,  the  point 
of  union  is  as  strong  as  any  other  part  of  the  bar.  Precaution 
must  be  taken  to  keep  the  surfaces  clean  and  free  from  scales, 
which  would  interfere  with  perfect  welding  contact  and  are  so 
apt  to  form  in  a  thin  fire  of  the  forge.  Scales  are  due  to  the 
oxidation  of  the  iron,  which  while  red-hot  is  not  sufficiently  sur- 
rounded by  ignited  carbon  to  consume  the  free  oxygen  of  the  air. 
When  the  layer  of  fuel  is  thin,  or  where  too  much  blast  is  given, 
the  nascent  iron  absorbs  the  oxygen.  Once  formed,  the  scales 
cannot  be  melted  or  fused  off,  but  their  formation  may  be  pre- 
vented by  a  liberal  covering  of  fuel  over  the  iron  or  by  sprinkling 
borax  over  the  surfaces  during  the  heat.  Sometimes  sand  may 
do  instead  of  borax,  though  it  requires  a  higher  temperature  for 
its  fusion.  With  proper  regulation  of  fire,  steel  may  also  be 
welded  to  iron  with  advantage;  as,  for  example,  the  steel  ends 
of  a  pick,  which  receive  the  wear  of  the  rock,  are  welded  into  the 
iron  head  of  the  pick-eye. 

Steel  is  a  compound  of  carbon  with  iron  in  varying  propor- 
tions. H.  M.  Howe,  in  his  "Metallurgy  of  Steel,"  says  that  steel, 


MINERS'   TOOLS.  631 

in  its  specific  sense,  is  "a  compound  of  iron  possessing  or  capable 
of  possessing  decided  hardness  simultaneously  with  a  valuable 
degree  of  toughness  when  hot  or  when  cold,  or  both.  It  includes, 
primarily,  compounds  of  iron  combined  with  from,  say,  0.3  to 
2  per  cent  carbon,  which  can  be  rendered  decidedly  soft  and  tough 
or  intensely  hard  by  slow  and  rapid  cooling,  respectively;  and, 
secondarily,  compounds  of  iron  with  chromium,  tungsten,  man- 
ganese, titanium,  and  other  elementary  compounds,  which,  like 
carbon  steel,  possess  intense  hardness  with  decided  toughness." 
"The  terms  'iron'  and  'steel'  are  employed  so  ambiguously 
and  inconsiderately  that  it  is  to-day  impossible  to  arrange  all 
varieties  under  a  simple  classification."  The  various  adjectives 
qualifying  the  term  "spring/'  "shear,"  etc.,  apply  to  the  uses 
to  which  the  steel  is  put,  and  imply  a  certain  percentage  of  car- 
bon constituency. 

Hardening  Steel. — The  homogeneity  of  steel  and  the  presence 
of  carbon  imparts  to  it  a  capability  of  hardening  and  tempering 
to  a  degree  depending  on  the  temperatures  of  the  heating  and 
the  subsequent  cooling.  As  the  amount  of  carbon  increases, 
the  melting-point  of  the  iron  decreases;  and  this  greater  fusi- 
bility reduces  its  welding  quality. 

A  steel  is  said  to  be  "hardened"  if,  when  red-hot,  it  is  suddenly 
cooled.  The  reason  for  this  change  is  not  readily  understood, 
though  it  is  in  some  degree  owing  to  the  presence  of  the  carbon; 
for  pure  malleable  iron  is  not  in  the  least  affected  by  the  opera- 
tion, whiile  both  steel  and  cast  iron  are  hardened  to  a  marked 
degree.  That  attained  by  cast  iron  by  plunging  it  into  a  cooling 
fluid  is  not  so  great  as  with  steel.  The  greater  the  difference 
in  temperature  between  the  steel  and  the  cooling  fluid  and  the 
shorter  the  time  of  hardening,  the  harder  it  becomes. 

The  fluids  used  for  the  cooling  bath,  are  water,  oil,  mercury, 
and  lead,  according  to  the  care  to  be  exercised  in  the  process. 
Water,  having  the  highest  specific  heat,  performs  the  operation 
of  cooling  most  quickly.  The  steel,  however,  scales  somewhat 
and  becomes  excessively  brittle.  The  oil  is  of  slower  action 
anu  is  believed  also  to  supply  from  its  decomposition  some  carbon, 


632  MANUAL  OF  MIXING. 

which  carburizes  with  the  steel.     Lead  and  mercury  are  used 
in  large  establishments. 

Tempering. — This  process  follows  hardening.  The  use  to 
which  the  tool  is  to  be  put  determines  the  degree  of  hardness 
desired,  but  the  latter  cannot  be  secured  accurately  without  skill 
in  one  operation  when  water  is  used.  It  may  be  attained  by  a 
bath  of  molten  lead  for  certain  uses,  but  usually  the  steel  is  sub- 
jected to  a  second  operation  of  annealing. 

The  hardened  steel  is  again  heated  to  a  red  heat  and  thus  is 
softened  slightly.  Its  point  is  then  plunged  for  an  instant  into 
a  cooling  fluid  to  harden  the  lower  end.  Withdrawing  the  article 
and  rubbing  off  the  scales,  the  heat  from  above  the  end  will  be 
conducted  to  the  edge.  A  series  of  colors  plays  over  the  surface 
in  succession  as  the  temperature  increases.  Beginning  \vith  a 
light  straw,  passing  through  the  shades  of  yellow,  brown,  purple, 
and  blue,  the  effects  of  the  chill  are  partially  removed,  till,  if  the 
end  should  become  red,  all  its  hardness  will  have  been  removed. 
When  the  steel  point  assumes  the  color  desired,  leaving  a  given 
degree  of  hardness,  the  tool  is  finally  cooled  by  immersion.  A 
tool  plunged  at  a  straw  is  very  hard,  while  one  allowed  to  anneal 
by  reheating  to  a  blue  is  quite  soft. 

Caution  is  urged  that  the  plunged  tool  while  tempering  be 
not  held  too  long  a  time  at  a  certain  color-line,  for  it  has  a  strong 
tendency  to  break  there  when  in  use.  The  tool  should  be  slightly 
waved  in  the  water.  Pieces  which  are  to  be  tempered  throughout 
must  be  allowed  to  "soak";  i.e.,  become  uniformly  hot  before 
plunging. 

The  Correct  Temper  for  Tools. — The  proper  color  for  a 
given  ground  is  only  ascertained  by  experience.  Generally 
speaking,  the  picks  and  drills  are  stopped  at  a  straw  if  intended 
for  hard  rock  and  carried  nearly  to  a  blue  for  mild  ground.  It 
is  always  desirable  to  preserve  the  toughness  of  the  steel  as  far 
as  possible;  therefore  the  lowest  color  is  selected  which  is  com- 
patible with  the  sendee  to  be  performed.  A  high-carbon  steel 
is  given  a  lighter  color  than  steel  of  low  carbon. 

Metal- working   tools   are   given  a  pale  straw-yellow;   wood- 


MINERS'    TOOLS.  633 

working  tools,  a  brownish  tint;  hatchets,  saws,  etc.,  a  light  purple; 
picks,  to  a  rose;  cold-chisels,  to  an  orange-rose;  keydrifts,  orange; 
rock-drills,  yellow-orange;  screw-cutting  dies,  light  yellow,  and 
hammer  faces,  a  pale  straw.  Few,  if  any,  miners'  tools  are  carried 
to  a  blue  temper.  They  would  be  too  soft.  Crowbars  should 
never  be  hardened,  as  they  would  be  too  brittle  under  severe 
strain. 

Bits  which  have  been  properly  tempered  will  wear  down 
uniformly  till  too  blunt  for  further  service.  Those  which  are  too 
brittle  will  be  found  to  have  broken  or  cracked  off.  The  former 
might  be  again  tempered  at  a  lighter  color  and  the  latter  should 
have  been  softened  further  during  the  sharpening. 

REFERENCES. 

The  Testing  of  Explosives  in  Belgium,  Coll.  Guard.,  Jan.  9,  1903;  The 
Theory  of  Misfires  and  some  Conclusions  of  Practical  Value,  E.  H.  Weis- 
kopf,  Jour.  Chem.  and  Met.  Soc.  of  S.  Africa,  Nov.  1902;  Theory  of 
Misfires,  Coll.  Guard.,  Vol.  LXXXIV,  689;  The  Use  of  Explosives  in 
Mines,  Coll.  Guard.,  Vol.  LXXX,  738;  Explosives  in  Building,  Coll.  Guard., 
Vol.  LXXIX,  32;  Safety  Explosives,  Coll.  Guard.,  Vol.  LXXIX,  832. 

Shot-firing,  M.  &  M.,  Vol.  XX,  409,  426;  Blasting  Agencies,  M.  &  M., 
Vol.  XXII,  301;  Blasting  Rules,  M.  &  M.,  Vol.  XXII,  501;  Testing  Blast- 
ing Caps,  M.  &  M.,  Vol.  XXIV,  302;  The  Testing  of  Explosives  in  Bel- 
gium, Coll.  Guard.,  Jan.  9,  1903;  The  Theory  of  Misfires  and  some  Con- 
clusions of  Practical  Value,  E.  H.  Weiskopf,  Jour.  Chem.  and  Met.  Soc.  of 
S.  Africa,  Nov.  1902;  Modern  Explosives,  J.  S.  S.  Erame,  Nature,  April  5, 
1900;  The  Question  of  Underground  Magazines  for  Explosives,  Franz 
Prispisle,  Oesterr.  Zeitschr.  f.  Berg.-  u.  Hiittenswesen,  March  17  and  24,  1900; 
The  Theory  of  Safety  Explosives,  H.  Heise,  Gluckauf,  March  24,  1900; 
Explosives  on  Witwatersrand,  T.  Lane  Carter,  Eng.  &  Min.  Jour.,  Nov.  7, 
1903;  Testing  Mine  Explosives,  A.  W.  Warwick,  Mines  &  Min.,  Oct.  1902; 
Some  Tests  of  the  Relative  Strength  of  Nitroglycerine  and  Other  Explo- 
sives, F.  N.  Clark,  Vol.  XVIII,  515;  Dynamite  Accidents  and  Preventions, 
James  Ashworth,  Coll.  Guard.,  Nov.  1896,  927;  List  of  Flameless  Explo- 
sives, Copy  of  Orders,  Coll.  Guard.,  June  1897, 1 129;  Preventions  of  Accidents 
from  High  Explosives,  James  Ashworth,  Coll.  Guard.,  Nov.  1896,  927;, 
Smokeless  Powder,  E.  &  M.  Jour.,  Vol.  LVI,  117;  Testing  Explosives  for- 
Coal-mines,  E.  &  M.  Jour.,  Vol.  LXI,  567;  Manufacture  and  Use  of.  Dyna- 
mite, Harry  Lee,  E.  &  M.  Jour.,  Vol.  LXI,  182. 


CHAPTER  VII. 

CHANNELERS,  DRILLS,  AND   COAL-CUTTERS. 

Machine  Rock-cutters. — The  successful  substitution  of  ma- 
chinery for  hand  labor  has  proven  a  most  important  advance  in 
mining  engineering.  Hard  rock  is  no  longer  an  obstacle  to 
extraction  of  fuel,  ore,  or  rock,  and  very  long  and  large  tunnels 
are  now  rendered  possible  in  a  comparatively  short  time.  The 
opening  of  mines  is  accomplished  in  such  short  time  and  with 
more  prompt  returns  for  the  investment  that  machine  drills  are 
•employed  to  the  exclusion  of  hand  labor  for  driving  shafts  and 
tunnels.  Every  form  of  hand-labor  tool  has  been  successfully 
imitated  and  extensively  introduced.  The  quarry  methods  of 
lewising  the  jumper,  saw,  chisel,  pick,  and  auger  find  their  counter- 
parts in  the  channeler,  percussive  drill,  coal-cutter,  and  diamond- 
drill. 

Quarrying. — The  quarrying  of  dimension  stone  was  formerly 
accomplished  by  the  trenching  along  lines  decided  upon.  Carried 
often  to  a  lo-foot  depth  and  wide  enough  for  a  man  to  operate 
his  pick,  these  trenches  wasted  much  good  material.  These 
trenches  are  now  replaced  by  channelers  and  gadders,  which 
dig  as  deep  as  desired,  but  only  2  or  3  inches  wide.  These 
machines  are  mounted  in  different  styles  and  cut  perfectly  true 
lines  at  any  angle  with  or  across  the  strata. 

Channelers. — For  extensive  quarries  these  machines  are 
mounted  on  a  portable  slining  carriage,  with  boiler,  rails,  etc., 
and  a  feed  which  automatically  moves  it  with  the  progress  of 
its  channel.  A  set  (gang)  of  five  cutters  receives  a  reciprocating 

634 


CHANNELERS,    DRILLS,   AND  COAL-CUTTERS.  635 


FIG.  298. — A  Single-gang  Channeler. 


636 


MANUAL  OF  MINING. 


motion  from  a  steam-piston,  through  a  connecting-rod,  or  through 
some  yielding  contrivance  from  the  cross-head  of  the  engine. 
The  latter  gives  an  elastic  blow  to  the  cutters.  Automatic  con- 
trivances keep  the  cutters  to  their  work.  Machines  are  also 
supplied  for  cutting  two  channels  at  a  desired  distance 
apart;  these  are  known  as  "double-gang  machines"  and  cost 


FIG.  299. — A  Broach-bit. 

from  $1200  to  $2000  complete.  With  three  men  and  400  Ibs.  of 
coal,  at  150  strokes  per  minute,  they  cut  from  75  to  400  sq.  ft. 
of  stone — the  former  in  marble,  the  latter  in  soft  lime — and  replace 
fifty  men  (Fig.  298). 

The  Broach-bit. — Many  quarries  employ  a  small  machine 
similar  to  the  channeler  mounted  on  a  transversing  and  long 
stout  bar  which  maintains  the  alignment  of  the  work.  Its  frame 
(Fig.  299)  is  comparatively  light  and  is  adjustable  to  a  high  or 
low  position.  Its  bit  may  be  a  broad  channeler  bit  and  drive 
rectangular  holes  in  any  direction  and  to  any  desired  depth,  or,  by 
permitting  the  rotation  of  the  bit  with  each  blow,  the  ordinary 


CHANNELERS,  DRILLS,  AND   COAL-CUTTERS.  637 

round  hole  may  be  drilled  by  the  same  machine.  The  broach- 
bit  proper  is  a  channeler-bit  which  breaks  down  the  partitions 
between  the  holes  drilled  by  a  previous  bit,  as  in  Fig.  299.  By 
this  machine  the  channeler  is  cut  to  any  length  and  depth  desired, 
the  line  between  the  holes  being  directed  by  the  stout  bar  of  the 
machine.  This  machine  is  also  employed  for  drilling  holes  for 
the  plugger  and  feather  work.  By  it  300  linear  feet  of  2 -foot 
holes  are  put  in  ten  hours,  or  70  square  feet  of  channeling  in 
sandstone  and  28  square  feet  in  granite. 

Lewising  is  accomplished  by  machine  by  fitting  a  tripod  with 
a  slot  movement  for  the  drill  body  such  that  three  parallel  holes 
can  be  drilled  with  a  single  setting. 

Power  Drills. — Reciprocating  drills  employing  steam  or  air 
as  a  motor  fluid  are  of  common  use  for  piercing  the  rock.  The 
construction  of  the  drill  is  the  same  for  either  fluid,  though  the 
name  given  to  the  drill  varies  with  the  fluid  used.  A  cylinder 
of  cast  iron  is  capable  of  sliding  in  a  guide  bed-plate  (Fig.  300) 
which  is  mounted  on  a  tripod  or  a  column.  The  cutting-tool 
is  clamped  to  an  extension  of  the  piston-rod  and  the  balance  of 
the  mechanism  is  such  as  is  necessary  to  obtain  a  return  of  the 
drill  with  its  progress  and  reversal  of  the  motion  of  the  piston. 
The  essentials  for  a  successful  drill  are  simplicity  in  construction; 
as  few  moving  parts  as  possible;  a  strength  commensurate  with 
light  weight  and  easy  transportation ;  variable  stroke  to  the  piston, 
whose  blow  shall  be  cushioned;  and  an  automatic  return  and 
revolution  of  the  drill.  The  weight  of  the  machine  must  be 
sufficient  to  receive  securely  a  shock  of  the  blow.  Thus  a 
piston  having  a  diameter  of  2^  inches  operating  under  an 
air  pressure  of  60  Ibs.  per  square  inch  would  receive  a  con- 
stant accelerating  force  of  294  Ibs.  with  every  stroke.  To 
counteract  this  recoil  its  mass  must  exceed  this  in  amount.  As 
the  steam-  or  air- drill  has  its  largest  sphere  of  usefulness  in 
remote  mining  districts  it  should  be  durable  and  require  a 
minimum  of  repairs. 

The  motor  fluid  underground  is  usually  compressed  air  and 
on  the  surface  steam,  with  an  ordinary  pressure  of  50  to  80  Ibs. 


638 


MANUAL  OF  MINING. 


per  square  inch.  The  horse-power  of  the  drill  is  estimated  as  a 
simple  steam-engine,  with  the  important  difference  that  the  ratio 
of  the  area  of  piston-rod  to  piston  is  larger.  Again,  the  steam- 


FIG.  300. — A  Steam-  or  Air-drill. 

engine  does  its  work  throughout  the  entire  stroke,  but  the  drill- 
engine  only  at  the  end  of  its  stroke.  Hence  it  can  never  work 
expansively.  The  air  enters  the  cylinder  and  propels  the  piston 


CHANNELERS,  DRILLS,  AND  COAL-CUTTERS.  639 

to  the  end  of  its  stroke,  when  the  attached  drill  strikes  the  rock. 
At  that  moment  the  piston  automatically  reverses  its  own  valve, 
which  admits  air  at  the  lower  end  of  the  cylinder,  while  a  ratchet 
and  spiral  device  slightly  turn  the  tool,  which  is  being  drawn  back 
for  the  next  blow.  As  the  work  to  be  done  on  the  return-stroke 
is  merely  to  lift  the  tool,  the  annular  area  of  the  piston  is  much 
less  than  that  on  the  other  side,  and  little  power  is  consumed. 
At  the  proper  point  in  the  up-stroke  the  valves  are  again  reversed 
and  the  operation  repeated. 

The  Automatic  Steam-valve. — The  main  difference  between 
the  drills  upon  the  market  is  in  construction  of  the  valve  and  the 
means  by  which  it  gives  a  reversal  of  motion  to  the  piston.  It  is 
desirable  for  rapid  penetration  that  the  number  of  blows  per 
minute  should  be  a  maximum,  but  structural  difficulties  place  a 
limit  to  the  length  of  time  between  each  blow.  It  is  essential  that 
the  full  air  pressure  be  exerted  at  the  instant  of  the  impact  and 
that  the  valve  should  not  reverse  until  the  blow  has  been  struck. 
Then  it  should  reverse  instantly,  permitting  only  sufficient  steam 
to  remain  in  the  clearance  space  under  the  piston  to  safely  cushion 
the  blow  against  the  front  head  of  the  cylinder.  The  average 
speed  is  about  200  blows  per  minute.  The  frequency  of  the 
blow  varies  with  the  ability  of  the  machine  and  is  altered  to 
suit  the  hardness  of  the  rock.  A  short  stroke,  light  blow,  and 
rapid  rate  give  the  best  progress  in  hard  rock,  and  a  hard  blow 
is  best  in  soft  rock,  provided  the  drill  does  not  "stick"  in  the 
hole. 

There  are  two  systems  of  moving  the  valves.  The  first 
requires  reversing  rod  and  tappets,  while  the  second  is  a  duplex 
system  requiring  a  fluid  to  reverse  the  motion. 

The  Tappet-valve. — In  the  earlier  forms  of  drills  the  valve 
was  operated  by  means  of  an  external  rod  with  an  exposed  three- 
armed  tappet  moved  by  a  projection  on  the  piston-rod,  as  in  the 
steam-pump.  The  rate  of  speed  of  a  drill,  however,  was  too 
high  for  the  mechanism  to  withstand  the  numerous  and  violent 
shocks  ensuing  at  the  time  of  the  reversal  of  the  valve. 


MANUAL  OF  MINING. 


m 


1 


The  later  forms  of  the  tappet  are 
concealed.  The  arc  of  its  motion 
is  reduced  and  a  more  compact 
machine  is  the  result  (Fig.  301). 

The  valve  motion  was  positive, 
and  though  there  are  many  disad- 
vantages inseparately  connected  with 
it,  it  is  safer  in  the  hands  of  unskilled 
labor.  This  accounts  for  its  reten- 
tion in  our  present  successful  forms. 

The  little  Rand  giant  drill  (Fig. 
300)  is  a  tappet-moved  valve  and 
requires  less  dead  space  and  con- 
sumes  less  steam  than  the  Burleigh 
pattern,  which  it  resembles,  whose 
piston  operates  two  rockers  to  oscil- 
late  its  valve.  Centrally  in  the  cylinder 
a  three-toed  rocker  is  located,  the 
upper  toe  being  fitted  into  a  recess 
in  the  valve  which  it  moves  and  the 
two  lower  toes  being  alternately  rocked 
by  the  piston,  to  produce  the  reversal 
of  motion.  By  thus  separating  the 
spindles  from  the  valves  and  the 
tappets  which  they  connect,  a  greater 
durability  was  obtained  than  in  the 
earlier  types  of  drills.  In  the  Ser- 
geant tappet-drill  the  valve  and  the 
rocker  are  also  in  a  single  three-toed 
piece,  driven  alternately  by  either 
end  of  the  piston. 

In  the  Sergeant  (Fig.  303)  the 
piston-valve  is  moved  by  exhaust 
steam  from  the  opposite  ends.  An 
auxiliary  slide-valve  moves  over  the 
arc  of  a  circle  by  shoulders  on  the 


CHANNELERS,  DRILLS,  AND  COAL-CUTTERS.  641 

piston,  opens  and  closes  the  ports,  and  is  a  trigger  regulating 
the  movement  of  the  main  valve.  There  are  no  openings  in 
the  side  of  the  cylinder  and  no  ports  for  the  piston  to  close;  the 
exhaust  remains  open  at  one  end  till  the  blow  is  struck,  when  the 
valve  reverses  immediately. 

The  Fluid-driven  Valve. — This  is  a  piston-valve  which  reverses 
the  motion  of  the  piston  by  opening  the  appropriate  ports  for 
admission  and  exhaust  of  steam  or  compressed  air,  being  itself 
cause  to  oscillate  in  its  valve- chest  by  the  motor  fluid  which  is 
automatically  admitted  to  one  head  or  the  other  of  its  piston. 
It  is  adopted  in  the  Rand  "Slugger"  (Fig.  302),  the  Ingersoll 
"Eclipse"  (Fig.  303),  the  Schram  drill,  and  others.  The  first 
two  named  are  of  similar  construction  and  action,  and  are  shown 
in  section.  Two  port- holes  connect  the  annular  groove  in  the 
piston  with  each  opposite  end  of  the  valve-chest  and  are  opened 
or  closed  by  the  piston  passing  over  them;  the  supply  for  one 
end  and  the  exhaust  to  the  other  end  of  the  valve-chest  are 
simultaneously  opened.  The  annular  groove,  therefore,  is  a  gen- 
eral exhaust  outlet  for  the  valve  steam,  while  the  motor  steam 
is  exhausted  by  the  valve  connecting  the  inlet  passage  with  the 
exhaust-pipe. 

There  are  no  means  on  the  piston  in  Fig.  302  for  maintain- 
ing a  steam-tight  or  air-tight  separation  of  the  two  ends.  When 
wear  occurs  the  steam  pressure  is  lost  and  leakage  ensues;  the 
exhaust  becomes  imperfect  and  the  valves  may  fail  to  act  properly. 
When  in  practice  the  machine  is  found  not  to  reciprocate  properly 
the  fault  is  usually  found  to  be  due  to  wear. 

The  Schram  Drill. — An  account  of  the  Schram  and  the  Darling- 
ton drills  is  to  be  found  in  Andre's  "Mining  Machinery,"  from 
which  the  following  is  taken :  "  Schram's  consists  of  a  slide-valve 
and  a  slide-rod  that  admit  steam  to  the  cylinder  for  raising  the 
piston  and  drill.  When  the  piston  passes  a  certain  front  port- 
hole, steam  enters  through  it  into  the  back  of  the  valve-chest 
at  the  same  time  that  the  front  valve-chest,  through  the  other 
port  and  the  hollow  circular  groove  of  the  piston,  communicates 
with  the  exhaust-pipe.  Steam  then  works  full  pressure  on  the 


642 


MANUAL  OF   MINING. 


CHANNELERS,  DRILLS,  AND   COAL-CUTTERS.  643 

slide  cylindrical  rod,  which,  with  the  slide-valve,  is  forced  towards 
the  front  valve-chest,  so  that  the  back  steam  passage  is  open 
to  the  cylinder,  and  the  front  steam  passage  connects  with  the 
exhaust-pipe.  The  piston  moves  forward,  and,  when  it  passes 
the  back  port,  allows  the  steam  to  enter  the  front  valve-chest 
at  the  same  time  that  the  back  valve-chest,  through  its  back  port 
and  the  circular  groove  of  the  piston,  communicates  with  the 
exhaust.  The  slide-rod  is  forced  back,  the  front  steam  passage 
opens,  and  the  back  passage  communicates  v/ith  the  exhaust. 
The  slide  is  in  the  form  of  two  spindle-valves,  so  that  it  remains 
in  position  without  recoil,  and  the  annular  groove  of  the  piston 
is  always  in  communication  with  the  exhaust. 

"  The  Darlington  Drill  has  only  two  working  parts, — an  extreme 
of  simplicity:  a  cylinder  and  its  cover,  and  a  piston  and  its  rod. 
The  piston  is  made  to  operate  as  a  valve.  The  inlet-pipe,  having 
open  connection  with  the  cylinder,  always  furnishes  the  pressure 
to  lift  the  drill,  which  rises  whenever  there  is  no  pressure  on 
the  back.  On  its  way  up,  the  piston  first  covers  the  exhaust 
(above  the  inlet),  and  then  uncovers  an  equilibrium  passage, 
by  means  of  which  communication  is  established  between  the 
front  and  back  ends  of  the  cylinder.  Then  air  or  steam  enters 
and  operates  over  the  greater  area,  at  the  back,  and  first  checks 
the  upward  movement,  soon  overcomes  it,  and  finally  produces 
a  forward  motion.  The  propelling  force,  now,  is  dependent 
upon  the  difference  of  area  between  the  back  and  front  of  the 
piston.  On  its  way  down  it  soon  cuts  off  the  equilibrium  passage 
and  the  air  can  only  enter  at  the  inlet;  the  steam  operates  by 
expansion  for  a  short  space,  till  the  piston  has  passed  and  uncovered 
the  exhaust-port,  when  a  discharge  takes  place  as  the  blow  is 
being  struck.  One  fact  is  noticeable,  that  the  amount  of  steam 
used  is  only  that  necessary  for  the  down  stroke;  for  that  used 
to  raise  the  drill  escapes  by  the  equilibrium  passage  to  the  top." 

The  Drill-rods.  —  The  steel  drill  tool  is  of  a  diameter  depend- 
ing on  the  size  of  piston,  and  from  f  to  i^  inches,  according 
to  the  intensity  of  the  blow  to  be  struck.  The  smallest  size 
is  attached  to  a  2 -inch  piston  and  the  largest  size  mentioned 


644  MANUAL  OF  MINING. 

to  a  5-inch  piston,  corresponding  to  blows  of  200  Ibs.  and  1200 
Ibs.  respectively.  The  average  size  of  the  mining  drill  is  a  3-inch 
piston  with  i-  or  i^-inch  steel.  The  drill-rod  is  clamped  by  a 
heavy  split  chuck  locking  into  the  enlarged  end  of  the  piston-rod, 
and  reciprocates  with  it.  The  end  of  the  drill  or  shank  varies 
in  dimensions  according  to  the  piston  size.  For  a  2-inch-diameter 
piston  the  shank  is  f"X-g",  for  a  2-inch  piston  i$"X7",  for  a 
piston  of  4^-inch  diameter  and  proportionate  sizes  for  the  inter- 
mediate diameters.  It  reciprocates  through  a  split  or  solid  spool 
front  head,  which  is  provided  with  a  stuffing-box  and  a  gland 
to  prevent  the  escape  of  steam.  The  latter  head  is  adapted  to 
air-power,  while  the  split  fronts  are  preferred  for  steam-drills. 

The  bit  is  usually  forged  with  a  bluff  cutter  edge  for  the 
strength  and  is  provided  with  the  usual  flare.  For  rocks  which 
do  not  crush,  but  chip  in  large  fragments,  a  sharper  edge  will  do 
better  execution.  The  cutter  is  dressed  by  special  tools  to  the 
X,  I,  Z,  or  S  form,  each  having  its  specific  value,  the  S  form 
being  more  likely  to  maintain  a  round  hole. 

A  set  of  tools  accompanying  each  drill  is  graded  according 
to  the  amount  of  feed  which  the  drill  is  capable  of.  The  shortest 
and  the  stoutest  drill  of  the  set  is  about  15  inches  in  length  and 
the  other  lengths  are  successively  longer  by  the  amount  of  feed 
of  the  machine.  The  longest  drill  in  the  set  corresponds  to 
the  maximum  depth  of  the  hole.  A  machine  feeding  15  inches 
and  drilling  a  hole  10  feet  deep  will  have  8  drills  in  a  set,  the 
difference  between  their  lengths  being  15  inches.  For  a  5-foot 
hole,  3  or  4  drills  constitute  the  set.  The  bits  should  be  hardened 
in  such  a  manner  as  to  have  durability  corresponding  to  the 
length  of  the  feed.  A  bit  usually  will  require  sharpening  after 
4  inches  of  progress. 

During  drilling  the  tool  must  be  turned  uniformly,  as  other- 
wise the  hole  may  become  rifled.  By  rifling  is  understood  the 
tendency  of  a  drill  to  cut  a  triangular  hole  instead  of  circular 
hole,  which  it  strikes  successively  at  one  point  or  along  a  given 
line.  With  the  straight-edge  cutter  the  tendency  to  rifling  is 
stronger  than  with  any  other  form. '  In  rock  of  an  unhomogeneous 


CHANNELERS,  DRILLS,  AND  COAL-CUTTERS.  645 

character  rifling  is  common.  With  deep  holes  this  may  also 
result  in  a  deflection  of  the  hole  from  a  straight  line. 

To  prevent  rifling  and  to  produce  rotation  of  the  drill  through 
a  small  arc  during  each  stroke,  the  fluted  bar-nut  constituting 
the  long  thread  ratchet   is   a 
usual     appliance.        In     the 
Burleigh   and  the  Darlington 
drills   the   device   is   a   spiral 
feather  on  the  piston-rod,  re- 
cessed   into   a    grooved   piece 
in   the   cylinder-head.       It   is 
toothed  and  held  by  a  detent,  FlG-  304--The  Rifle-bar. 

which  permits  it  to  turn  on 

the  forward  stroke,  but  prevents  turning  during  the  up  stroke 
of  the  engine.  In  the  Ingersoll  drill  a  grooved  bar  fitting  into 
the  back  of  the  piston  turns  it  on  the  back  stroke  and  is  itself 
allowed  to  rotate  on  the  down  stroke  (Fig.  304). 

In  the  Darlington  drill  the  ratchet  turns  the  piston  and  drill 
on  the  up  stroke,  and  itself  turns  during  the  down  stroke. 

In  the  Schram  pattern  an  auxiliary  piston  turns  the  drill,  the 
former  being  driven  by  the  fluid  in  a  manner  similar  to  that  of  the 
valve. 

The  Feed.— Whatever  may  be  the  pattern  of  the  drill,  the 
method  of  feeding  the  cylinder  and  its  piston  to  the  work  is  the 
same.  It  is  always  under  control  of  the  drill-runner  and  is  capable 
of  adjustment  to  the  progress  in  the  rock.  The  cylinder  and 
its  tool  is  allowed  to  move  in  a  guideway  which  is  rigidly 
mounted  in  the  frame  by  the  means  of  a  square  thread  feed-screw 
which  is  turned  by  a  crank  at  the  rear  end  of  the  machine.  The 
drill-runner  usually  stands  at  work  with  one  hand  on  the  feed- 
handle  crank  and  the  other  on  the  steam-throttle  valve. 

Cushioning  the  Blow. — The  maximum  percussive  effect  is 
obtained  when  the  blow  is  struck  under  the  full  head  of  steam 
or  air.  Such  a  dead  blow  is  highly  desirable;  but  on  account 
of  the  shock  to  the  machine,  for  the  subsequent  repairs  this  is 
unadvisable.  A  clearance  space  is  provided  in  which  is  enclosed 


646 


MANUAL   OF  MINING. 


a  small  amount  of  air  to  act  as  the  cushion  or  elastic  buffers  are 
inserted  to  terminate  the  stroke.  The  former  is  less  expensive 
than  the  latter. 

The  Drill  Supports. — A  rigid  support  is  a  rigid  adjunct  to 
the  drill,  and  several  types  of  mountings  are  provided,  each 
having  a  special  end  in  view,  though  a  machine  can  be  shifted 
from  one  style  to  another.  In  tunnels  and  shafts  where  the 
ranges  of  holes  have  approximately  parallel  directions,  it  is 
clamped  directly  upon  a  stout,  hollow,  cylindrical  column  (Fig. 
305),  or  upon  an  arm  projecting  from  it. 


Ho.  305. — The  Column  Support  for  Drills. 

This  admits  of  drilling  several  holes  from  one  position  of 
support.  Jack-screws  at  one  end  clamp  the  column,  while 
claws  at  the  other  end  bear  against  blocks  resting  upon  the  rock. 
The  column  can  be  had  in  various  lengths.  It  weighs  about 
30  Ibs.  per  foot.  When  placed  horizontally,  as  in  shafts  (Fig. 
305),  it  is  known  as  a  shaft-bar.  [When  used  in  the  vertical 
position  it  is  spoken  of  as  a  column. 

The  tripod  is  the  more  common  drill  support,  having  a  wider 
range  of  position  than  the  column  because  of  the  universal  joint 


CHANNELERS,  DRILLS,  AND  COAL-CUTTERS.  647 

at  its  head  (Fig.  306).  Its  stability  depends  upon  its  weight, 
and  additional  dead  weights  are,  therefore,  clamped  on  each 
leg  as  in  Fig.  300. 


FIG.  306. — Blocking  out  Dimension  Stone. 

The  drill  has  no  intricate  mechanism  and  can  be  intrusted 
even  to  unskilled  labor.  Care  should  be  taken  in  beginning 
operations  that  the  drill  strikes  squarely  upon  the  face  of  the  rock, 
and  with  short,  light  blows.  It  can  do  more  work  with  the  con^ 
sumption  of  less  powrder,  steel  and  blacksmith  wrork  than  can  hand- 
drills.  It  can  be  set  up  in  any  place  capable  of  accommodating 
a  double-hand  gang  of  men,  and  gives  little  trouble  in  placing 
into  position.  It  cannot  be  adopted  in  small  veins  or  where  the 
mineral  is  friable  or  occurs  in  thin  streaks. 

The  comparative  merits  of  the  several  types  of  machines  can- 
not be  stated,  for  in  one  mining-camp  the  Rand  drill,  and  in 
another  the  Ingersoll  drill,  are  exclusively  used.  The  Waring 
drill,  the  National,  and  the  Burleigh,  are  employed  for  the  same 
operations  in  the  different  mines  of  the  same  camp  with  appar- 
ently equal  success.  One  range  prefers  the  little  Giant,  but  in 


648  MANUAL  OF  MINING. 

another  mine  it  is  discarded  for  Slugger  pattern  of  the  same 
company.  In  like  manner  a  preference  may  be  displayed  for 
the  Eclipse-  or  Sergeant- drill  made  by  the  same  company.  They 
are  all  highly  commended,  and  their  improvement  in  the  inland 
community  may  be  a  matter  of  accident  or  of  natural  selection 
after  periods  of  test. 

The  fluid- moved  valve- drill  style  seems  the  favorite  pattern 
for  hard  rock;  but  whether  under  all  circumstances  it  is  the 
best,  one  would  not  dare  to  aver.  The  nature  of  the  rock,  the 
proper  air  pressure,  the  rate  of  speed,  and  the  proportion  of 
rotary  motion  necessary  for  a  maximum  effect  vary  so  largely  that 
a  comparison  of  the  numerous  published  results  cannot  be  made. 

Progress  and  Cost  by  Machine-drills. — -An  average  of  nine 
neighboring  mines  in  conglomerated  rock  shows  machine-driving 
and  stoping  to  be,  respectively,  22  and  36  per  cent  cheaper  than 
hand,  and  sinking  4  per  cent  dearer,  with  the  progress  respectively 
60,  54,  and  38  per  cent  more  rapid.  The  gain  in  time  during 
sinking  more  than  compensates  for  the  cost.  In  iron- mines  the 
product  of  the  machine  labor  costs  about  one  fourth  as  much 
as  that  of  hand  labor.  The  consumption  of  fuel  or  air  per  drill 
may  be  calculated  in  the  same  manner  as  for  the  ordinary  steam- 
engine.  The  cost  of  the  average  mining  drill  is  about  $325, 
and  a  complete  plane  with  its  drills  and  suitable  compressor,  etc., 
is  $7000. 

Systems  of  Machine-drilling.  —  With  the  advance  made  in 
the  use  of  the  power-drill,  the  systems  of  drilling  and  blasting 
rock  have  undergone  corresponding  changes.  In  hand-work  the 
object  sought  when  placing  a  hole  for  drilling  and  blasting  is 
as  much  to  secure  a  good  bench  for  the  succeeding  shot  as  to 
break  a  maximum  of  ground.  With  machine-drilling,  simul- 
taneous shooting  is  practised  and  numerous  holes  are  drilled 
over  the  entire  area  of  shaft  or  drift,  and  fired  in  volleys  with- 
out regard  to  the  probable  condition  of  the  face  after  the  blast. 
The  removal  and  the  resetting  of  the  machine  after  each  blast 
occupy  so  much  time  that  a  maximum  number  of  holes  is  drilled 
with  each  set-up,  and  the  time  thus  saved  more  than  compensates 


CHANNELERS,  DRILLS,  AND  COAL-CUTTERS.  649 

for  the  value  of  the  powder  wasted  by  not  conforming  to  the 
fundamental  principles  of  blasting,  as  in  single  shooting. 


Systems  of  Drilling  Holes. — There  are  two  general  systems  in 
practice  for  the  drilling  of  holes,  the  first  being  known  as  the 
centre-cut  system,  employed  for  tunnels  and  shafts  of  an  area 


650  MANUAL  OF  MINING. 

larger  than  fX$f.  The  second  system  is  known  as  the  Brain 
radial  system,  which  is  employed  in  headings  only  large  enough 
to  accommodate  one  machine.  In  either  system  as  many  machines 
as  possible  are  set  up  against  the  face.  Three  are  employed  in  a 
single-track  tunnel;  two  in  a  n-foot  heading;  and  as  many  as 
six  are  arranged  in  a  double- track  tunnel  27  feet  wide.  It  is  not 
unusual  to  mount  two  drills  on  a  single  column.  The  holes 
are  drilled  as  determined  upon,  and  after  loading  are  blasted 
for  an  advance  of  6  to  10  feet  with  each  set-up. 

The  Centre-cut  System  is  of  almost  universal  acceptance  in 
America.  Vertical  tiers  of  holes  are  drilled  over  the  face,  as  in 
Fig.  307. 

The  method  of  placing  the  holes   is  shown  in  Fig.  308,  the 

arrows   indicating   approximately   the    direction   in    which   they 

are  drilled.     The  top  or  back  holes 

o>  a  .  . 

|  aa   are    pointed   a   little    upward   in 

t<*  order  to  obtain  the  top  height  and  to 

b\     °^^     |    ^^»     Jb  break   down   the   rock  left   after   the 

_c     j  other    holes    have    been    fired.     The 

side    holes    bb    are    pointed    a    little 

•      •       *"       "*       *  upward   and   outward  to  obtain    the 

b\      ^*       ^-^.c     tb  full  width  of  tunnel  after  the  centre 

FIG.   3o8.-Direction   of   Holes   holes    are    fired'     The    Centre    tiers   of 
for  the  Centre-cut  System  for  holes  cc  are    pointed  centreward,  the 

Wide  Gangways.  *  . 

bottom  ones  of  the  tier  being  given  an 

upward  slant,  the  top  ones  a  downward  slant.  Sometimes  a 
plunger- hole  d  is  drilled  and  fired  with  the  centre  holes.  The  holes 
are  fired  in  rounds.  Those  at  the  centre,  having  the  hardest 
work,  are  charged  with  a  higher  grade  of  powder  and  throw  out 
a  wedged-shaped  mass;  the  side  tiers  b,  being  charged  with  a 
lower  grade  of  dynamite,  are  fired  later  together,  after  which  the 
back  and  trimming-holes  are  exploded  together  to  complete 
the  cut. 

In  the  use  of  the  centre-cut  system  for  wide  tunnels,  in  con- 
nection with  the  American  methods  of  tunnelling,  as  illustrated 
in  Figs.  309  and  264,  the  centre  holes  n  (Fig  310),  fired  first, 


CHANNELERS,  DRILLS,  AND-COAL  CUTTERS. 


05* 


facilitate  the  work  of  the  next  two  rows  2  2,  after  which  the  holes 
3  3  are  fired  in  volleys,  succeeded  by  the  trimming-holes  4  4. 

The  profile  should  be  finished  with  the  advance  of  the  face, 
regularly  checked  up  by  the  surveyors,  for  it  seems  difficult  to 
maintain  the  proper  floor  level,  the  tendency  invariably  being 
to  gradually  raise  the  floor  wiih  the  progress  of  the  work  in  an 
unconscious  effort  to  insure  drainage. 


FlG.  309. — Drilling  Bench-holes  in  the  American  System  of  Tunnelling. 

In  driving  such  a  tunnel  the  work  is  conducted  in  three  shifts 
of  eight  hours  each.  During  the  first  shift  the  drill-runners  and 
their  helpers  are  engaged  in  operating  the  machines  and  drilling 
the  holes  to  their  proper  depths.  The  powder-men  follow;  they 
remove  the  machines  from  the  face  to  a  safe  distance,  and  charge 
the  holes  with  explosive  and  ignite  them.  In  the  third  shift  the 
broken  material  is  removed  by  loaders,  who  tram  the  car  sfrom 
the  face  to  the  point  of  discharge  and  also  deliver  the  drill  and 
appliances  to  the  face  for  the  drillers. 


6S2 


MANUAL  OF  MINING. 


Drilling  in  Benches.— The  bench  of  the  tunnels  may  be 
a. tacked  in  two  sections,  as  in  Fig.  309,  A  and  B,  or  in  one  only, 
as  in  Fig.  310.  In  the  former,  two  wall-holes,  one  or  two  trans- 
verse rows  of  four  top  holes  downward,  and  half  a  dozen  bottom 
holes,  lift  each  bench  with  every  other  shift.  Fifty-four  feet  of 
progress  a  week  is  the  record  on  a  very  hard  sandstone  bench 
I4'X27'.  This  bench- work  is  accomplished,  not  only  more 


FIG.  310. — The  Centre-cut  System  of  Drilling  Wide  Tunnels. 

rapidly,  but  also  with  a  powder  consumption  per  cubic  yard  of 
rock  of  about  one  half  that  in  the  heading  above  it. 

The  Centre-cut  System  in  Shafts. — The  method  of  placing 
holes  in  shafts  does  not  materially  differ  from  that  in  the  head- 
ings. In  Fig.  311  is  the  elevation  of  a  6-foot  cut  in  a  shaft  whose 
area  is  io'X2o'.  The  centre  holes  are  placed  10  feet  apart 
approaching,  but  not  intersecting.  Six  pairs  of  holes,  spaced 
2  feet  apart,  extend  across  under  the  shaft.  The  holes  bb  are 
in  two  pairs  of  rows  of  five  each,  the  first  being  given  a  greater 
inclination  than  the  second  row.  They  are  fired  in  separate  tiers 
of  ten  holes  each.  Occasionally  all  of  the  twenty  holes  are  fired 


CHANNELERS,  DRILLS,. AND  COAL-CUTTERS, 


653 


together,  followed  by  the  squaring-up  holes  cc.  For  a  deeper 
cut  the  centre  holes  (Fig.  312)  are  in  two  tiers,  those  in  tier  a  nearly 
meeting  at  6  feet  from  the  shaft  floor  and  the  other  bb  at  u  feet. 
The  other  holes  are  driven  as  before.  In  this  case  the  6-foot 
centre  is  opened  first,  followed  in  sections  to  the  sides  by  the 
others. 


X 

x  & 

X.    K 

a  x  x  . 

b       x. 

x 

x    y 

x  x 

&    X 

*       b: 

x  xa 

xx  x 

FIG.  311.— The  Six-foot  Centre  Cut 
for  Shafts. 


FIG.  312. — The  Ten-foot  Centre  Cut 
for  Shaft-sinking. 


Brain's  Radial  System. — This  is  designed  for  small  faces 
which  can  accommodate  but  one  machine  at  a  time.  Like  the 
centre-cut  system,  it  is  equally  applicable  to  shafts.  All  the 
holes  to  be  drilled  are  put  from  one  position  of  the  machine,  which 
is  removed  only  when  firing  is  to  be  done.  This  minimizes  the 
time  lost  in  shifting.  In  a  certain  case  the  machine,  from  a  position 
4  feet  8  inches  from  the  bottom,  2  feet  from  the  top,  and  2  feet 
6  inches  back  from  the  face,  put  twenty-nine  holes  with  a  total 
length  70  feet,  advancing  3  feet  with  an  average  of  2.4  cubic  feet 
broken  rock  per  lineal  foot  of  hole. 

The  holes  are  comparatively  shallow,  though  they  vary  greatly 
in  length,  those  making  the  smallest  angle  with  the  face  being 
the  longest.  Four  ranges  of  holes  are  drilled.  Sometimes  a  few 
extra  squaring-up  and  lifting  holes  are  necessary  to  trim  the 


654  MANUAL  OF  MINING. 

periphery  of  the  drift,  but  ordinarily  the  firing  of  the  most  angling 
holes  first  breaks  out  the  rock  to  daylight  and  opens  a  face  for 
the  other  successive  rounds.  The  advance  cannot  be  large, 
for  neither  deep  nor  angling  holes  are  possible  in  a  narrow  drift. 
In  a  drift  8  feet  wide,  two  settings  of  the  machine  are  sometimes 
made  drilling  from  near  each  wall  and  thus  forming  a  modified 
centre-cut  plan.  In  some  mines  a  practice  prevails  of  cutting 
a  horizontal  range  of  bottom  holes,  two  ranges  of  holes  looking 
downward,  and  a  top  row  to  break  out  horizontal  instead  of 
vertical  wedges;  this  plan  requires  a  bar- mounting  for  the  drill, 
in  a  drift  of  say  fx&. 

The  Continuous  Process  of  Diamond-drilling. — Gen.  Henry 
Pleasant's  method  of  shaft-sinking  is  a  novel  and  eminently  suc- 
cessful application  of  the  diamond-drill.  One  or  more  diamond 
drilling-machines  are  set  up  over  the  site  of  the  shaft,  and  bore 
vertical  holes  as  deep  as  the  shaft  is  to  be  carried.  The  machines 
are  moved  to  new  positions  and  additional  long  holes  bored. 
The  operation  is  continued  until  the  entire  area  of  the  shaft  is 
pierced  by  holes  at  suitable  distances  apart.  One  shaft  had 
thirty-five  holes  bored  over  its  area  of  25'  8"Xi3'  10"  to  a  depth 
of  300  feet  in  six  weeks  by  three  machines. 

When  the  "continuous  process"  is  completed,  the  machines 
are  removed  for  the  blasting.  The  holes  are  filled  with  sand  or 
water  for  the  full  length,  except  in  the  upper  3  or  4  feet,  wrhich 
are  treated  like  short  holes,  charged  with  dualin  and  fired, — the 
central  ones  first.  When  the  debris  has  been  cleared  away,  the 
shaft  will  have  advanced  3  or  4  feet.  A  few  feet  of  each  hole 
are  again  cleaned  out,  loaded,  and  fired.  Thus  each  section 
advances  with  an  uninterrupted  alternation  of  shooting  and 
hoisting.  Though  it  is  not  always  cheaper  per  cubic  foot,  it 
effects  a  great  saving  in  time,  and  quick  access  under  ground 
may  prove  the  element  essential  to  the  success  of  the  undertaking. 

Boring  Headings.— The  diamond-drill  cannot  be  used  advan- 
tageously for  tunnel  work  or  driving  headings.  Percussion-drills 
are  there  used  almost  exclusively,  though  rotary  machines  have 
been  employed  for  boring  out  a  heading  almost  to  full  size.  The 


CHANNELERS,  DRILLS,  AND   COAL-CUTTERS.         .    655 

Stanley  header  represents  this  class  of  machine.  At  one  opera- 
tion a  series  of  cutters  on  a  rotating  boring-head  grinds  away 
the  whole  face  for  the  core  of  a  heading  some  7  feet  in  diameter. 
One  was  used  in  the  Mersey  subaqueous  tunnel.  It  travelled  at 
the  rate  of  39  inches  per  hour  and  executed  its  work  satisfac- 
torily in  the  argillaceous  chalk. 

The  approximate  cost  of  sinking  and  tunnelling  work  varies 
with  the  character  of  the  rock,  the  depth,  and  excavation,  of  dis- 
tance from  the  surface;  the  efficiency  of  the  workmen  also  deter- 
mines largely  the  economy.  In  the  New  house  tunnel,  with  an 
area  of  144  sq.  ft.,  the  cost  of  driving  was  $28.80  per  foot,  in  the 
Wooster  tunnel  it  was  about  $40  per  foot,  and  in  the  Aspen  tunnel 
$19  per  foot.  The  cost  of  driving  is  usually  greater  through 
soft  ground  than  through  hard  ground,  on  account  of  the  increased 
quantity  of  timbers  required  and  the  difficulties  of  draining  the 
water.  The  cost  of  driving  a  rock  tunnel  5'  X  f  on  contract 
without  timbering  is  $8  per  foot.  An  upraise  can  be  driven  at  a 
cost  of  between  $5  and  $10  per  foot,  varying  with  the  size  and 
the  character  of  the  timbering  needed.  A  two- compartment 
shaft  with  the  timbers  can  be  sunk  6'  X  12'  in  the  clear  at  an 
•expense  of  $25  per  foot. 

Electric  Rock-drills. — The  various  attempts  to  introduce  an 
electric  percussion-drill  to  replaced  compressed  air-drills  have 
resulted  in  only  a  partial  success.  The  advantages  which  elec- 
tricity present  in  transmission  over  compressed  air  are  sufficient 
to  suggest  an  installation  of  drills.  The  difficulty,  however,  lies 
in  the  inability  to  produce  a  reciprocating  motion  without  the 
use  of  solenoid. 

The  Solenoid  Drills. — Two  coils  of  wire  are  wound  in  the 
form  of  a  spiral  through  which  a  current  is  alternately  passed  in 
the  one  direction  and  the  reverse.  If  a  steel  core  is  passed  through 
the  adjacent  coils  and  a  current  of  electricity  be  passed  through 
one  of  them  the  coil  will  be  repelled  through  the  other  coil.  The 
latter  being  next  charged,  will  give  a  reverse  action  to  the  coil 
and  repel  it  toward  the  first.  This  sets  up  an  oscillatory  move- 
ment which,  if  the  machine  be  perfect  enough  and  the  coils  carry 


656  MANUAL  OF  MINING. 

sufficient  current,  will  accomplish  work  similar  to  the  percussive 
machine  (Fig.  313). 


ki^.  jij. — A  Solenoid  Drill.  FIG.  314. — A  Gardner  Drill. 

The  Siemens,  Halske,  and  Gardner  Electric  Drills  are  based 
upon  the  attempt  to  produce  indirectly  a  reciprocating  motion. 
The  current  does  not  enter  the  drills,  but  drives  gearing  housed 
in  the  drill  cylinder  by  means  of  a  motor  and  a  flexible  shaft. 
The  drill  is  mounted  on  an  adjustable  column  arm  or  on  a  tri- 
pod (Fig.  314).  This  machine  produces  a  rapid  succession  of 
light  blows  rather  than  a  small  number  of  heavy  blows.  It 
requires  2  to  3  horse-power  to  operate  it.  It  is  also  quite  heavy. 

The  Box  Electric  Drill  dispenses  with  the  flexible  shaft.  In 
the  rear  of  the  drill  is  a  motor,  geared  by  a  pinion  to  a  wheel 
on  the  crank-shaft.  At  the  other  end  of  the  crank  is  a  pinion 
which  operates  a  train  of  gearing,  turning  the  drill  after  each 
blow.  Between  the  crank-shaft  and  the  small  cylinder  is  a 
connecting-rod. 

Reciprocating  motion  is  imparted  to  the  cylinder  by  the 
crank.  The  heavy  piston  with  a  thick  piston-rod  constitutes 
the  hammer  whose  weight  determines  the  blow. 

Each  end  of  the  moving  cylinder  is  filled  with  air  at  atmos- 
pheric pressure,  which  being  alternately  compressed  and  rarified 
cushions  the  blow.  Two  ports  located  somewhat  centrallv  at  the 


CHANNELERS,  DRILLS,  AND  COAL-CUTl  ERS.  657 

side  of  the  moving  cylinder  are  provided  for  admitting  air  when- 
ever the  piston  uncovers  them.  The  weight  of  the  drill  with  its 
motor  is  about  350  Ibs.,  the  latter  alone  being  100  Ibs. 

The  Adams  All-steel  Drill  is  simple  in  mechanism,  has  few 
parts  easily  gotten  at,  and  all  wear  taken  up,  since  every 
bearing  is  adjustable.  The  drill  is  actuated  by  a  loose  rod  run- 
ning through  the  gear-case  and  motor,  and  this  imparts  motion 
through  a  pair  of  bevel  gears  and  crank-shaft  to  the  draw-bar 
and  piston,  the  latter  being  cushioned  with  gangs  of  helical 
springs  which  serve  to  absorb  the  shock  of  the  blow  to  protect 
the  parts  from  injury  and  also  to  store  up  energy  on  the  back 
stroke,  which  it  can  expend  in  forcing  the  steel  into  the  rock. 
These  springs  prove  extremely  useful  when  reaming  or  for 
pulling  the  drill  out  of  a  fissured  hole.  The  drill,  striking  from 
575  to  600  blows  per  minute,  will  consume  less  power  for  its 
work  than  an  air-drill.  It  consumes  three  horse-power  at  the 
generator  to  operate  it. 

It  is  claimed  that  in  point  of  first  cost  the  electric  drill  has 
the  advantage  over  an  air-drill,  for  a  one-drill  plant  complete, 
operated  electrically,  costs  one  half  of  an  air  compressor  and 
drill.  The  simple  and  smooth-running  dynamo  is  less  noisy 
than  the  complicated  air-compressor  as  a  source  of  power.  The 
objection  to  the  electric  drill,  that  it  provides  no  means  for  ven- 
tilating the  workings  as  does  an  air-drill,  is  met  by  the  equally 
valuable  advantage  that  light  can  be  obtained  from  the  same 
wire  that  supplies  the  power.  The  comparative  efficiencies  of 
the  rotary  electric  drills,  reciprocating  electric  drill  with  flexible 
shaft,  the  solenoid  reciprocating  drill,  and  the  air  reciprocating 
drill  are  respectively  as  i  :  1.66  :  4.5  :  10. 

Coal-cutting  Machines. — Coal  is  mined  by  machine  in  the 
same  manner  as  by  hand  labor.  Holes  may  be  bored  by 
machinery  and  the  powder  charges  in  them  exploded,  or  the 
coal  may  be  undercut  or  sheared  by  machine  and  broken  down 
with  or  without  the  aid  of  powder.  For  boring  holes  in  the  coal 
various  types  of  portable  augers  are  employed,  such  as  shown 
in  the  following  chapter.  For  the  undermining  of  coal  there 


658  MANUAL  OF  MINING. 

are  two  types  of  machines,  one  depending  on  percussion  and 
the  other  on  abrasion,  as  produced  by  the  chisel-edged  teeth. 
Shearing-machines  are  of  a  similar  type  of  construction  to  the 
undermining  machines.  The  percussive  machines  are  identical 
in  appearance  and  construction  to  the  air-  or  steam-drills  em- 
ployed in  rock  with  the  exception  of  the  character  of  the  bit. 
Of  abrasive  machines  there  are  three  classes  using  sharpened 
teeth  attached  around  a  bar  which  advances  broadside  on  the 
coal  face,  on  a  link  chain  advancing  normally  against  the  face, 
and  on  the  edge  of  a  rotary  wheel  at  the  side  of  the  machine 
advancing  parallel  to  its  face.  The  first  two  classes  are  breast 
machines  and  the  one  last  named  is  a  longwall  machine.  The 
motor  power  for  any  of  these  types  may  be  electricity  or  air. 

The  same  character  of  labor  that  the  miner  performs  Li 
removing  and  cutting  a  groove  is  imitated  by  the  undermining 
machines  in  the  floor  under  the  coal  or  on  the  bottom  layer  of 
the  coal  itself.  This  groove  is  carried  as  far  back  from  the  face 
as  possible  and  over  the  entire  width  of  the  room.  The  shearing- 
machine  cuts  a  vertical  groove  along  either  side  of  the  rib  into 
the  coal  face  and  to  the  depth  the  machine  can  reach,  beginning 
at  the  roof  and  extending  to  the  floor.  The  use  of  either  of  these 
machines  removes  from  the  miner's  toil  the  most  laborious  por- 
tion of  his  work  'as  well  as  the  most  hazardous  phase  of  his  occu  - 
pation.  The  normal  effort  of  the  digger  exerted  under  the 
unfavorable  conditions  existing  in  the  room  is  most  wasteful U 
applied  and  produces  in  addition  an  excessive  amount  of  fip<i 
coal,  which  is  lost.  The  machine  produces  more  coal  with  less 
waste  in  a  much  shorter  time.  It  reduces  the  labor  of  the  minei 
and  results  in  subdividing  the  work  formerly  imposed  upon 
one  man  and  increases  the  efficiency  of  each  man  by  specializing 
each  branch  of  the  work.  From  four  to  eight  loaders  follow  each 
machine  and  a  total  of  twelve  men,  on  the  average,  are  employed 
in  a  mine  for  each  machine  in  service.  Its  introduction  results 
in  an  increase  in  the  earning  capacity  of  each  miner  besides 
improving  the  conditions  under  which  his  labor  is  applied. 

The  introduction  of  machines  decreases  the  number  of  delays 


CHANNELERS,  DRILLS,  AND  COAL-CUTTERS.  659 

of  standing  shots.  It  gives  a  steadier  output  and  concentrates 
the  operations  of  the  mine  because  less  territory  must  be  kept 
open  for  the  desired  capacity.  It  requires,  however,  a  more 
systematic  development  of  the  mine. 

In  Illinois  the  Legg  machine  is  used  in  driving-rooms,  and 
elsewhere  the  Harrison,  Jeffrey,  Lechner,  and  Sullivan  machines. 
In  Europe  the  machines  in  vogue  are  known  as  the  Marshall 
and  Frith. 

The  Requirements  of  the  Machine. — It  should  be  light  and 
capable  of  being  handled  by  two  men  and  occupy  small  floor 
space  or  height,  to  admit  of  being  moved  around  and  between 
the  roof-props.  It  should  be  capable  of  starting  in  a  corner  of  a 
pillar  or  a  loose  end,  and  of  cutting  a  groove,  from  wall  to  wall 
of  the  room,  to  any  height,  right  handed  or  left  handed. 

The  Percussive  Machines. — The  reciprocating-drill  is  largely 
employed  for  underholing  the  coal  in  the  Mississippi  Valley 
States.  It  is  often  known  as  the  punch- drill,  and  has  as  its  chief 
representative  the  Harrison  machine  (Fig.  315),  which  requires 
little  explanation.  It  is  a  modification  of  the  compressed-air 
drill,  having  for  a  valve-motor  a  single  rotary  device.  It  is 
mounted  on  low  wheels  on  a  platform  inclined  toward  the  face 
to  enable  it  to  move  freely  while  making  a  cut;  it  is  guided  to  its 
work  by  handles  on  either  side  and  is  prevented  from  recoiling 
by  wooden  blocks  back  of  the  wheels.  In  addition  to  the  drill- 
runner,  one  helper  is  necessary  to  remove  the  chippings  from 
the  face  of  the  work.  An  air  pressure  of  about  60  Ibs.  to  the 
square  inch  is  used  expansively,  the  point  of  cut-off  being  fixed 
by  the  runner  according  to  the  hardness  of  the  coal.  About  200 
strokes  are  delivered  per  minute.  Usually  three  bits  are  supplied 
with  each  machine  in  lengths  of  2,  4,  and  6  feet.  A  considerable 
slack  coal  is  produced  while  digging  a  deep  undercut  of  5  or 
6  feet,  its  height  being  16  or  18  inches  at  the  face  and  3  inches 
at  the  rear.  The  lateral  dimensions  and  shape  of  the  groove 
are  determined  by  the  reach  of  the  machine,  but  the  entire  face 
of  a  room  is  entirely  undercut  by  a  succession  of  such  grooves,  for 
which  purpose  the  platform  and  machine  are  shiftsd  sidewise. 


66o 


MANUAL  OF  MINING. 


CHANNELERS,  DRILLS,  AND  COAL-CUTTERS.  66l 

The  average  work  of  the  machine  is  80  lineal  feet  of  face  in  ten 
hours,  which  is  equivalent  to  from  50  to  100  tons  of  coal  per  hour, 
according  to  the  height  of  the  seam.  The  machine  is  compact 
and  can  be  employed  where  the  roof  is  weak  and  the  timbers  are 
close  to  the  face  of  attack.  Its  weight  is  from  500  to  700  Ibs. 
and  its  height  rarely  exceeds  17  inches. 

The  Sergeant  Drill  Company  has  an  acceptable  adaptation 
of  its  drill  for  coal- work  which  is  largely  used  in  the  Southern 
States. 

A  variety  of  pick-machine  is  in  use  by  which  a  75-lb.  pick 
is  swung  by  a  bell- crank  lever  at  a  weight  of  70  blows  per 
minute,  cutting  a  hundred  square  feet  of  ground  2  inches  high 
and  42  inches  deep. 

The  Breast-machine. — In  this  type  the  coal-cutter  with  its 
motor  is  mounted  on  a  frame  which  is  caused  to  slide,  inside  of  a 
substantial  base  frame,  toward  the  face  or  breast  of  the  coal,  and 
under  it  to  the  limit  of  the  slide.  The  cutter  consists  of  a  num- 
ber of  hardened  steel  bits  on  an  edge  of  a  wheel  or  over  a  rotating- 
bar  in  front  of  the  machine ;  or  it  may  be  a  series  of  bits  inserted 
into  an  endless  link-belt.  The  wheel  or  bar  is  forced  forward, 
cutting  a  width  of  grooves  24  to  44  inches  to  a  depth  of  about 
6  feet.  The  groove  is  from  2\  to  5  inches  high.  The  chain  is 
drawn  horizontally  and  uniformly  around  four  wheels  at  the 
corners  of  the  sliding-frame,  and  as  its  cutters  pass  in  front  of 
the  machine  the  coal  is  cut  to  about  the  same  width  and  depth 
as  above  mentioned.  After  each  cut  the  motor  and  its  cutter 
are  withdrawn,  the  standard  frame  is  moved  along  the  face  a  dis- 
tance equal  to  the  width  of  its  groove,  and  from  the  new  position 
the  next  advance  is  made.  The  power  employed  may  be  either 
air  or  electricity,  the  former  being  preferred,  unless  the  mine 
contains  considerable  gas. 

In  Fig.  316  is  shown  the  Jeffrey  bar-cutter.  This  is  a  bar 
fitted  with  numerous  cutting-chisels,  rotated  by  a  link-chain 
shown  at  the  side  of  the  machine.  The  bar-cutters  are  kept  to 
their  work  by  a  feed  at  the  rear;  their  frame  is  shown  advanced 
slightly  from  its  starting  position.  A  pair  of  jacks,  one  at  the 


662  MANUAL  OF  MINING. 

front  and  another  at  the  rear,  braced  by  a  screw  against  roof 
and  coal,  gives  stability  to  the  machine.  Moreover,  the  jack 
at  the  front  in  a  measure  assists  in  holding  up  the  coal  from  bear- 
ing excessively  upon  the  cutter-rod.  This  machine  is  still  ser- 
viceable in  some  coals,  though  its  cutters  have  a  tendency  to 
climb  up  into  the  coal.  This  is  a  great  annoyance  to  the  operator. 
It  also  acts  in  its  abrasion  across  the  bedding-plane  of  the  coal 
instead  of  parallel  to  it,  as  does  the  chain-machine,  and  in  this 
respect  is  not  so  efficient. 

The  more  modern  type  of  breast- machine  manufactured  by 
the  Jeffrey  Machine   Company  is  illustrated  in  Fig.  317.     This 


pIG    3I7> — A  Jeffrey  Breast-machine. 

chain-cutter  is  the  logical  outcome  of  the  bar- cutter  of  the  past. 
An  endless  chain  carries  a  number  of  bits  made  of  f-inch  steel, 
having  a  chisel  cutting  edge  f  inch  wide.  The  chain,  revolved 
by  the  motor  and  suitable  gearing,  is  advanced  automatically. 
The  bits  project  from  the  chain  sufficiently  to  make  a  cut  4$ 
inches  on  either  side  of  it  (Fig.  318).  The  machine  weighs  about 
2000  Ibs.,  but  is  easily  removed  to  a  new  position  by  releasing 
the  jacks  and  skidding  the  frame,  after  which  the  jacks  are  reset. 
The  voltage  of  the  electric  current  is  500,  and  35  amperes  are 
consumed.  Excellent  results  are  obtained  in  gaseous  coal,  splint 
coal,  or  block  coal. 


CHANNELERS,  DRILLS,  AND  COAL-CUTTERS. 


663 


Longwall  Machines. — By  inserting  the  chisel-bits  into  the 
circumference  of  a  rotary  disc  which  is  operated  at  high  speed  by 
the  electric  motor,  projecting  it  from  the  side  of  the  machine, 


a  longwall  cutter  is  produced  by  which  the  machine  may  be  caused 
to  move  parallel  to  the  face  of  the  coal,  while  the  wheel  or  toothed 
disc  undercuts  the  coal  for  a  depth  equal  to  a  little  over  half  its 
diameter. 


C64 


MANUAL  OF  MINING. 


The  Winstanley  machine  consists  of  a  rotary  toothed  disc, 
capable  of  being  tucked  away  under  the  carriage  or  turned  out 
against  the  face  and  revolved  by  two  oscillating  cylinders.  With 
an  air  pressure  of  30  Ibs.  per  sq.  in.  and  a  machine  weighing 
1500  Ibs.,  mounted  on  a  carriage  moving  along  the  track,  the 
progress  of  the  cutter  is  70  sq.  ft.  of  underholing  in  an  hour. 

The  longwall  machine  placed  on  the  market  by  the  Jeffrey 
Machine  Company  is  represented  by  Fig.  319.  This  is  a 
one-rail  machine,  balanced  in  such  a  manner  that  the  cutter- 
wheel  may  pass  over  all  irregularities  in  the  floor.  A  series  of 


FIG.  319. — A  Longwall  Machine. 

gears  are  arranged,  winding  a  rope  on  a  drum  to  move  the  machine 
along  the  rail  at  a  rate  fixed  by  the  operator  between  8  and  25 
inches  per  minute.  The  cutter-wheels  are  from  3  to  6  feet  in 
diameter,  according  to  the  thickness  of  the  coal-bed.  The  bits 
inserted  on  the  circumference  of  the  wheels  make  a  cut  of  about 
4  inches  wide.  On  the  top  of  the  cutter-wheel  are  gear-teeth 
engaging  with  the  pinion  and  the  engine-shaft,  by  which  the 
rate  of  rotation  of  the  wheel  can  be  regulated  to  that  which  is 
most  suitable  for  the  coal.  This  machine  is  said  to  undercut  an 
average  of  600  lineal  feet  of  coal-face  per  day. 

Shearing  Coal-cutters. — The  shearing  of  coal  by  machine  may 
be   accomplished  by   a  percussive   machine,   mounted   on  high 


CHANNELERS,  DRILLS,  AND  COAL  CUTTERS.  665 

wheels  or  by  a  chain-cutter  turned  in  position  to  a  vertical  plane. 
The  reciprocating-machine,  when  shearing  coal,  is  mounted  on 
wheels  of  a  diameter  of  3  feet,  or  even  more,  enabling  it  to  assume 
a  wide  range  of  position  in  making  the  vertical  cut  from  a  high 
angle  pointing  upward  to  a  steep  angle  pointing  downward. 

One  variety  of  the  shearing-chain  cutter  is  shown  in  Fig.  320. 
It  is  mounted  on  four  posts,  each  supplied  with  a  jack-screw,  to 
secure  it  to  the  roof  and  floor  with  suitable  clamps  that  permit 
of  a  rotary  movement  about  the  column  and  an  adjustment  to 
any  height.  The  armature-shaft  of  the  motor  is  parallel  to 
the  centre  rail.  The  cut  is  commenced  at  the  top  of  the  coal 
and  a  groove  as  wide  as  the  machine  will  allow  and  as  deep  as 
it  will  reach  is  finished,  after  which  it  is  lowered  for  successive 


FIG.  320. — A  Coal-shearing  Machine. 

and  adjoining  cuts.  The  cutters  travel  downward  along  the 
coal,  thus  always  drawing  the  chippings  down  and  out. 

Many  of  these  coal-cutters  are  made  adjustable  in  position 
for  undercutting  or  shearing.  The  latter  is  not  an  economical 
method  of  procedure  as  compared  with  undercutting,  but  is 
serviceable  in  firm  coals  and  gaseous  mines  where  it  is  advisable 
to  use  powder. 

Another  shearing-machine,  made  by  the  "Sullivan  Machine 
Company,"  is  of  the  pick  type,  mounted  on  a  truck  and  pivoted 
to  swing  up  and  down  to  the  point.  It"  usefulness,  however, 
is  confined  to  the  centre  of  headings  and  it  cannot  be  employed 
for  shearing  ribs.  The  machine  makes  a  cut  6  inches  wide 
and  from  7  to  8  feet  deep,  averaging  30  lineal  feet  in  a  ten- 
hour  shift. 


666  MANUAL  OF  MINIXG. 

The  Comparative  Advantages  of  Coal-cutting  Machines. — 
Coal-cutting  machines  can  be  employed  to  advantage  underground 
when  the  coal  is  firm  and  in  seams  which  are  of  standard  dimensions, 
free  from  eccentricities,  and  under  a  strong  roof  with  a  good  floor. 
They  are  employed  when  it  is  necessary  to  develop  the  colliery 
rapidly  and  where  it  is  essential  to  obtain  a  maximum  proportion 
of  large  coal.  Th?y  have  an  advantage  over  handwork  in  concen- 
trating the  field  of  operations.  They  require  less  men  and  are 
especially  advantageous  where  shooting  off  the  solid  rock  is 
practised  and  where  there  is  no  soft  underlying  layer  in  which 
the  miner  can  underhole  the  coal.  They  cannot  be  employed 
to  advantage  in  wet  mines,  where  the  work  is  intermittent,  where 
the  pressure  of  the  roof  bears  too  heavy  on  the  coal  face,  or  where 
the  layer  under  the  coal  is  pyritiferous,  or  in  the  anthracite  coal, 
which  is  full  of  slip- joints. 

The  statistics  furnish  ample  evidence  that  the  powder  con- 
sumption in  machine  mines  is  less  per  ton  of  coal  than  in  hand 
mines.  Likewise  the  number  of  fatal  and  serious  accidents 
occurring  in  machine  mines  is  much  less.  It  usually  requires 
from  three  to  five  additional  machines  to  maintain  an  uninter- 
rupted work  for  seven  machines. 

A  machine  can  cut  a  groove  of  6  feet  depth  and  44  inches 
width  in  from  3^  to  5  minutes,  and  with  the  time  occupied  in  making 
six  changes  the  undercutting  of  a  room  20  feet  wide  can  be  effected 
in  two  hours'  time.  Assuming  thirty  minutes  as  necessary  to 
shift  the  machine  to  the  'next  room  and  to  reset  it,  it  will  be 
possible  to  underhole  four  rooms  in  a  day  writh  the  employment 
of  two  men.  This  in  a  standard  seam  of  4  feet  thickness  corre- 
sponds to  144  tons.  By  handwork  with  two  men  at  a  face,  eleven 
rooms  would  have  to  be  kept  open  to  equal  the  supply  of  this 
one  machine,  which  operates  in  four  rooms.  A  mine  producing 
looo  tons  of  screened  coal,  1300  tons  of  run-of-mine  coal,  can 
obtain  it  by  nine  machines  in  thirty-six  rooms  or  by  200 
miners  in  TOO  rooms.  The  operations  therefore  are  more 
concentrated,  as  less  territory  is  kept  open  than  in  the  hand- 
work, and  this  is  an  important  feature  when  there  is  taken 


CHANNELERS,  DRILLS,  AND  COAL-CUTTERS.  667 

into   account  the  maintenance  of  roads,  ventilation,  and  supply 
of  timber. 

One  marked  objection  to  coal-cutting  machines  lies  in  the 
quality  of  coal  which  is  produced.  This  arises  from  the  fact  that 
the  coal  is  broken  in  large  lumps,  which  contain  an  undue  pro- 
portion of  impurities.  They  require  less  explosive,  and  indeed 
necessitate  the  use  of  light  charges;  since  heavy  shot-firing  weakens 
the  pillars  and  wastes  the  coal. "  Cutting-machines  require  broad 
rooms  to  obtain  the  maximum  quantity  of  coal  at  the  cheapest 
rate,  and  this  results  in  the  reduction  in  the  size  of  the  pillars, 
with  the  consequent  sacrifice  of  much  coal  in  the  ribs  and  stumps. 
Moreover  its  accumulation  in  the  gob  tends  to  develop  spontane- 
ous combustion. 

Percussion  Machines  versus  Chain  Machines. — The  former 
class  of  undercutting  machine  is  especially  advantageous  over 
the  other  types  when  the  roof  is  so  weak  as  to  require  timbering 
close  to  the  face  and  where  the  bearing- in  layer  under  the  coal 
is  too  stony  or  too  full  of  pyrites  to  be  cut  with  continuous-motion 
cutters.  It  has  also  an  advantage  where  there  is  excessive  pres- 
sure from  the  roof  upon  the  face  of  coal.  It  reduces  the  cost  of 
mining  to  quite  a  considerable  degree  even  when  the  cost  of 
operation,  interest,  and  maintenance  are  included.  It  is,  however, 
slow  and  produces  a  great  deal  of  slack  coal,  besides  doing  less 
work  in  a  given  time  than  can  the  breast  machines.  The 
latter  can  undercut  in  fire-clay  or  in  soft  under  layer  quicker 
than  the  punching-machine.  In  fire-clay  the  chain  machine 
is  preferred,  while  for  soft  coal  or  slate  the  stronger  disc  or  wheel 
must  be  used. 

The  breast  and  longwall  machines,  employing  chains,  wheels, 
or  rotating  bars,  can  be  operated  by  electricity  or  by  air,  whereas 
the  reciprocating  machine  must  be  operated  by  compressed  air. 
The  cost  of  installing  their  power  plants  is  about  the  same  for 
an  equipment  of  equal  capacity  and  a  distance  of  one  mile 
from  the  power-house.  For  greater  distances  electricity  ad- 
mitted is  far  cheaper  than  compressed  air.  In  comparing  the 
cost  accounts  of  producing  coal  by  either  type  of  machine  not 


663 


MANUAL  OF  MINING. 


only  are  wages,  materials,  interest,  and  depreciation  to  be  con- 
sidered, but  also  the  profit  on  the  amount  of  coal  lost  by  each 
method. 

Comparison  of  Hand-mining  with  Machine- work.  —  The 
amount  of  slack  dirt  and  waste  which  must  be  allowed  for  in 
hand-mining  is  as  a  rule  50  per  cent  more  than  with  machine. 

co/v- 


P!G   221. — Showing  Position  of  Sullivan  Coal-cutter  in  a  Room. 

When  the  net  profit  it  represents  is  included  the  saving  of  machine 
coal  over  hand-mined  coal  is  still  greater.  The  machine-cut 
groove  6  feet  deep  saves  three  cubic  feet  of  coal  per  yard  of  face 
in  the  reduced  slack.  A  machine  cannot  be  employed  to  advan- 
tage in  seams  as  thin  as  those  which  can  be  mined  by  hand. 
Neither  can  the  breast  or  longwall  machine  be  employed  in  as 


CHANNELERS,  DRILLS,  AND  COAL-CUTTERS.  66) 

narrow  and  confined  quarters  as  can  a  percussive-machine  cut- 
ter. As  to  output  per  year  13,000  tons  may  be  cited  as  an  aver- 
age production  for  either  class  of  machine.  Hand  labor  as  com- 
pared with  machine  is  expensive,  and  the  tendency  therefore  is 
to  confine  the  former  to  narrow  work  and  the  extraction  of  coal 
from  the  ribs  and  pillars.  At  the  present  time  more  than  one 
quarter  of  the  soft-coal  product  of  the  United  States  is  being 
machined,  with  a  rapid  increase  in  its  adoption  with  time.  When 
hand  labor  can  no  longer  be  employed  profitably  the  ribs  and 
pillars  must  be  recovered  in  the  second  or  room  working  by  some 
other  type  of  machine  than  those  employed  at  the  present  time. 
However  rapid  the  present  machine  may  undercut  or  shear  the 
coal  it  is  not  sufficiently  rapid  for  the  extraction  of  the  rib  coal 
in  its  entirety  without  considerable  loss  and  risk.  The  longwall 
type  of  machine  will  probably  be  called  into  requisition  even 
though  it  may  differ  from  the  character  of  the  machine  employed 
in  the  rooms  and  workings  of  the  mine.  In  this  regard  the 
Sullivan  coal-cutter,  Fig.  318,  meets  the  requirement,  as  it  may 
be  moved  parallel  to  the  face  by  a  feed-chain,  as  shown  in 
Fig.  321. 

REFERENCES. 

Coal-mining  Machines,  R.  M.  Hazelton,  Ohio  Mine  Inspector,  1985,  22; 
Discussion  of  Present  Form  of  Machine  Cutters,  Cyrus  Robinson,  Amer. 
Mfr.,  1897,  121 ;  Efficiency  of  Modern  Mining  Machinery,  Cyrus  Robinson. 
Amer.  Mfr.,  1897,  588;  Machine  Mining  and  the  Labor  Question,  W.  E, 
Garforth,  Coll.  Guard.,  1897,  480;  Coal-cutting  by  Machine  in  Iowa,  Foster 
Bain,  Coll.  Guard.,  June  1897,  1085;  Coal-cutting  Machines  for  Pillar  and 
Stall  or  Narrow  Work  John  Davis,  Coll.  Guard.,  May  14,  1897,  918;  Coal- 
cutting  Machinery,  Chas.  Latham,  Coll.  Guard.,  1897,  133;  Notes  on  Coal- 
getting  by  Machinery,  T.  H.  Woodsworth,  Fed.  Inst.  M.  E.,  Vols.  VI  and 
VII;  Blakemore,  Vol.  XI,  179;  Machine  Mining  in  Ohio,  Amer.  Mfr.,  193. 

Relative  Advantages  of  Coal-cutters,  Anon.,  Coll.  Eng.,  Feb.  1897,  313; 
Coal-cutting  Machines  in  Longwall,  England,  T.  B.  A.  Clarke,  Coll.  Guard., 
Dec.  1896,  1078;  Electricity  in  Bituminous-coal  Mining,  Robert  M.  Hasel- 
tine,  Mine  Inspector,  Ohio,  1894,  18;  Coal-cutting  by  Machinery,  W.  Blake- 
more,  Trans,  of  the  N.  of  Eng.  Inst.  of  M.  &  M.  Eng.,  Vol.  XLV,  177;  Coal- 
cutting  Machinery,  John  B.  Atkinson,  Engineering  Association  of  the  South, 


6 70  MANUAL  OF  MINING. 

Pub.  No.  4;  Coal-mining  Machines,  A.  Dick,  Mineral  Industry,  Vol.  II, 
230. 

Electrical  Coal-cutting,  J.  T.  Burchell,  Eng.  News,  April  6,  1893,  334; 
Tunnels  in  Coal-mines:  Cost  of  Driving,  With  and  Without  Explosives, 
M.  Elce,  Coll.  Man.,  Dec.  1893,  222. 

Diamond  Drill  Prospecting,  J.  Parke  Channing,  Eng.  Mag.,  March 
1896;  Exploring  by  Diamond  Drill:  Cost,  etc.,  Archibald  Blue,  221;  4th 
Report,  1893,  164;  The  Diamond  Drill  for  Deep  Boring  Compared  with 
other  Systems  of  Boring,  Oswald  J.  Heinrich,  Amer.  Inst.  M.  E.,  Vol.  II, 
241;  Cost  and  Results  of  Geological  Explorations  with  the  Diamond  Drill 
in  the  Anthracite  Regions  of  Pennsylvania,  Lewis  A.  Riley,  Amer.  Inst. 
M.  E.,  Vol.  V,  303. 

Machine  Mining  and  the  Labor  Question,  W.  E.  Garforth,  Coll.  Guard., 
1897,  1085;  Coal-cutting  Machines  for  Pillar  and  Stall  or  Narrow  Work, 
John  Davis,  Coll.  Guard.,  May  14,  1897,  918;  Coal-cutting  by  Machinery 
in  Iowa,  Foster  Bain,  Coll.  Guard.,  June  1897;  Diamond  Drill  Prospecting, 
J.  Parke  Channing,  Eng.  Mag.,  March  1896;  Exploring  by  Diamond  Drill: 
Cost,  etc.,  Archibald  Blue,  Ontario  Bureau  of  Mines,  1895,  221;  Diamond 
Drill  Prospecting,  Rich.  A.  Parker,  S.  of  M.  Quart.,  Vol.  XVI,  31;  Curva- 
ture of  Diamond  Drill  Holes,  J.  Parke  Channing,  E.  &  M.  Jour.,  Vol.  LVIII, 
268;  Machine  Lining,  M.  &M.,  Vol.  XXII,  173;  Morgan  Gardner  Machine, 
M.  &M.,  Vol.  XXII,  237;  Longwall  Machines,  M.  &  M.,  Vol.  XXIV,  25; 
Pick  and  Shaft  Machines,  M.  &  M.,  Vol.  XXIII,  32  and  510. 

Coal-cutting  by  Machinery,  Blakesmore,  Trans.  Fed.  Inst.,  Vol.  XI,  179; 
Coal-cutting  Machines,  Jeffrey,  Horse-power,  Coll.  Eng.,  Nov.  1895,  75; 
Machines  in  Longwall,  England,  Coll.  Guard.,  Dec.  1896,  1078;  Coal- 
mining Machines,  Mineral  Industry,  Vol.  II,  230;  Coal-mining  and  Methods, 
Vol.  IV;  Electric  Coal-cutters,  Am.  Mfr.,  Jan.  i,  1897,  9;  Cutters  and  Drills 
in  Alaska,  Coll.  Guard.  Eng.,  Feb.  1897,  314;  Steam-shovel  in  Mining,  De- 
scriptives,  L.  Sup.  Mm.  Inst.,  Vol.  IV,  1896,  60;  Diagram  of  Forces,  Cas- 
sier's  Mag.,  Jan.  1895,  195;  In  Lake  Superior,  Coll.  Guard.,  June  18,  1897, 
1188;  Diamond  Drill  Prospecting,  Channing,  Eng.  Mag.,  March  1896; 
Stanley  Header  in  England,  Trans.  M.  &  M.  Eng.,  Vol.  XLV,  179. 


CHAPTER  VIII. 

BLASTING. 

The  Principles  of  Blasting. — The  principle  employed  in 
rupturing  rock  consists  in  subjecting  the  surface  of  a  subfacial 
cavity  regular  or  irregular  to  a  sudden  increase  of  pressure,  acting 
radially  outward.  When  the  agent  is  sufficiently  powerful  to 
produce  a  high  degree  of  compression  upon  the  surrounding 
rock  it  either  fractures  the  material  by  the  formation  of  a  con- 
geries of  crevices  or  it  shatters  it.  The  extent  of  the  destruc- 
tion depends  upon  the  intensity  of  the  pressure  and  the  cohesion 
or  the  toughness  of  the  material. 

Blasting-agents. — There  are  two  classes  of  blasting- agents. 
The  first  includes  all  varied  means  of  causing  a  prolonged  pres- 
sure of  greater  or  less  intensity,  while  the  second  develops  the 
force  more  or  less  instantaneously.  In  the  first  class  are  mechan- 
ical devices  such  as  the  steady  hydraulic  pressure  obtained  from 
a  ram;  that  of  pistons  actuated  by  compressed  air;  the  swelling 
of  slaking  lime;  and  the  spreading  produced  by  wedges.  The 
second  class  comprises  various  chemical  combinations  such  as 
black  powder,  nitroglycerine,  ammonite,  etc. 

Blasting-agents  of  the  First  Class.  —  These  are  employed 
for  driving  preparatory  works  in  the  fiery  mines  and  are  safer 
and  less  expensive  than  powder.  Lime  is  a  safe  effective  blasting- 
agent  in  coal,  and  can  be  made  up  into  cartridges,  which  after 
insertion  into  a  deep  hole  are  packed  tight  and  moistened.  In  a 
short  time  the  lime  slakes  and  expands  to  break  the  coal  away. 
Compressed  air  forced  into  a  cylindrical  case  by  a  powerful  air- 
pump  is  capable  of  exercising  such  a  pressure  as  may  be  sufficient 

671 


672 


MANUAL  OF  MINING. 


to  break  the  coal.  This  is  employed  in  some  coal  regions  and 
is  a  simple,  safe,  cheap  method.  Wedges  driven  by  any  of  these 
agents  are  similarly  employed  by  being  inserted  in  the  line  of 
the  cleavage  of  the  rock. 

Explosives. — An  explosive,  according  to  Andre,  is  a  mixture 
"capable  of  being  suddenly  transformed  into  gases  by  the  appli- 
cation of  heat."  In  this  sudden  evolution  of  gas,  in  a  space 
formerly  occupied  by  a  solid,  a  pressure  is  produced  upon  the 
confining  surface  to  the  volume  of  the  evolved  gas  to  that  of  the 
explosive.  The  expansive  force  of  the  gas  is  greater  as  the  tem- 
perature of  ignition  increases  and  with  the  rapidity  of  evolution  of 
the  gas.  If  the  latter  was  instantaneous,  the  maximum  pressure 
is  imparted  at  the  moment  of  explosion;  if  the  combustion  is  slow 
and  transmitted  from  grain  to  grain,  its  strength  is  dissipated 
over  a  longer  period  of  time,  and  the  pressure  is  less.  Thus 
the  strength  of  an  explosive  is  measured  by  its  specific  volume, 
the  amount  of  gas  it  produces,  the  temperature,  and  the  rapidity 
of  the  evolution. 

Below  is  a  table  of  the  relative  volumes  and  pressures  of  gas 
produced  by  the  perfect  combustion  of  i  Ib.  of  explosive  occupy- 
ing about  0.016  cubic  foot  of  volume. 


Volume  of  Gas. 

Heat-units. 

Relative  Force 

By  Single 
Explosion. 

^    By  • 

Detonation. 

Blasting-powder  

2.38  cu.  ft. 
5-09      " 

II.  01 

10.72      " 
9.76      " 

900,000 
570,000 
1,050,000 
1,240,000 
2,375»ooo 

I  .OO 
1.  08 
3-06 
3-15 

4-55 

4-34 
3.61 
6.46 
6.00 

10.00 

Chloride  of  nitrogen  

Guncotton  

Picric  acid  
Nitroglycerine  

When  an  explosive  is  ignited  the  heat  developed  in  its  com- 
bustion is  rapidly  communicated  from  grain  to  grain  and  the 
particles  decomposed  with  the  liberation  of  gases.  According 
as  these  two  phenomena  follow  each  other  slowly  or  quickly 
we  have  rending  or  shattering  powders,  of  which  the  representa- 


BLASTIXG.  673 

live  types  are  black  powder  and  nitroglycerine.  In  the  first 
the  gas  is  evolved  so  slowly  as  to  give  time  for  a  concentration 
of  pressure  along  a  line  or  lines  of  least  resistance.  This  is  the 
quality  desired  for  the  artilleryman  or  for  a  sporting  powder — 
ability  to  project.  The  slow  combustion  operates  upon  the 
small  mass  of  the  bullet,  which  can  instantly  take  up  a  very  high 
velocity  and  thereby  give  a  rapidly  increasing  space  for  the  evolved 
gases  to  escape.  If  the  bullet  or  plug  were  tight  and  moved 
slowly  it  would  burst  the  breech  or  muzzle. 

The  miner,  however,  desires  to  break,  and  this  property  is 
obtained  from  such  agents  as  rapidly  produce  gases  at  high 
initial  temperature  and  pressure.  Very  little  plugging  is  needed, 
for  the  concussion  produced  by  the  gases  of  the  quick  powder 
is  practically  instantaneous  and  the  wave  of  pressure  extends 
in  all  directions,  which,  being  resisted  by  the  rock,  spends  its 
force  there  and  shatters  it.  The  more  sudden  the  action  the 
more  local  is  the  effect. 

Igniting  and  Detonating  Explosives. — Without  attempting  to 
follow  the  history  of  blasting,  for  which  the  student  is  referred 
to  Rziha's  "Lehrbuch  dcr  Gesammten  Tunnelbaukunst,"  an 
enumeration  of  the  several  simple  and  compound  substances 
used  or  suggested  at  various  times  to  produce  concussion  may 
be  here  given  in  chronological  order:  Common  black  powder, 
picric  acid,  guncotton,  terchloride  of  nitrogen,  nitroglycerine, 
and  ammonite.  This  list  may  seem  brief,  but  a  longer  list  would 
be  merely  an  enumeration  of  the  varieties  obtained  by  the  sub- 
stitution of  a  single  constituent.  We  have  the  artillery,  sport- 
ing, and  blasting-powders,  composed  of  charcoal,  sulphur,  and 
saltpetre  in  varying  proportions;  picric  acid  and  picrates  with 
saltpetre  or  chlorate  of  potash;  guncotton,  combined  with  other 
explosives;  nitroglycerine,  with  admixture  of  absorbents  and 
dilutants.  The  result  is  that  we  have  various  grades  of  explosive 
compounds,  from  those  which  may  be  ignited  by  heating  to  a 
temperature  of  about  300°  C.  to  the  nitroglycerine,  which  requires 
a  shock.  In  other  words,  we  have  igniting  or  detonating  com- 
pounds. 


674  MAX  UAL  Ol'    MINING. 

Black  Powder. — The  term  black  powder,  or  gunpowder, 
embraces  mechanical  mixtures  of  carbon,  sulphur,  saltpetre 
varying  from  12,  8,  and  80  per  cent  respectively  for  sporting 
purposes  to  20,  16,  and  64  per  cent  for  open-air  blasting  and 
11.5,  17.5,  71  for  hard-rock  underground.  It  is  black  or  brown 
according  to  the  per  cent  of  carbon.  Its  constituents  are  pul- 
verized, compressed  into  a  cake,  granulated,  sifted,  glazed,  and 
dried.  The  size  of  the  grains  depends  upon  the  use  to  which  the 
povcder  is  to  be  put.  The  smaller- grained  powders  are  pre- 
ferred for  mining  purposes.  The  color  of  the  powder  varies 
from  black  to  dark  brown,  the  former  having  a  larger  percentage 
of  carbon  than  the  latter. 

Methods  of  Charging  the  Hole  with  Powder. — When  the  hole 
has  been  drilled  to  the  required  depth,  the  powder  is  either  poured 
into  it  from  the  can  or  is  inserted  as  a  cartridge.  The  cartridge 
is  a  tube  of  oiled  paper  which  is  closed  at  the  bottom  and  also 
at  the  top  after  the  black  powder  and  fuse  have  been  inserted. 

Black  powder  is  loaded  by  the  barrel  method  or  the  needle 
method.  In  the  latter  method,  which  is  more  common,  the 
cartridge  is  inserted  into  position  with  a  needle  projecting  from 
it  out  to  the  daylight.  Above  the  powder,  around  the  needle, 
and  filling  the  hole  for  some  distance  is  driven  a  very  soft-clay 
packing  material,  which  is  rammed  tight  into  position.  The 
needle  is  then  withdrawn  and  replaced  by  a  fuse  for  ignition. 
The  needle  is  a  slender-pointed  rod  of  copper  with  a  ring  at  the 
upper  end. 

In  the  barrel  method  the  powder  cartridge  is  pierced  by 
a  wire  which  leads  up  through  a  half- inch  copper  tubing,  or 
"barrel,"  that  extends  the  entire  length  of  the  hole.  Around  the 
barrel  is  tamped  the  packing,  after  which  the  wire  is  pulled  out 
and  replaced  by  a  fuse.  A  comparison  of  the  methods  shows 
a  preference  in  favor  of  the  barrel  because  less  and  poorer  quality 
of  tamping  may  be  used;  it  is  twice  as  fast;  the  cartridge  has 
less  opportunity  to  soak  water;  and  the  cheap  barrels  are 
recovered  after  shooting.  Some  anthracite  miners  use  a  f-inch 
\vrought-iron  pipe  for  the  blasting  bale. 


BLASTING.  675 

A   Blast-hole    Loader   on   the   Barrel   System  is  in  use  for 

loading  and  tamping  large  charges  of  powder  without  waste  or 
danger  of  premature  explosion.  Tubes  in  5 -foot  lengths  are 
screwed  together  with  a  funnel  at  the  top,  and  a  short  bronze 
discharge-tube  is  at  the  lower  end.  A  hickory  tamping-bar  com- 
pletes the  outfit.  A  system  so  complete  and  compact  saves  time 
and  powder  (Fig.  340). 

Tamping  and  Ignition  of  the  Powder. — To  confine  the  large 
volumes  of  gases  within  the  small  volume  occupied  by  the  powder 
a  tight  cover  of  melted  clay  is  rammed  on  top  of  the  charge  and 
packed  tight.  The  process  is  known  as  tamping. 

As  a  precaution  to  prevent  a  premature  explosion  the  mate- 
rial used  must  be  free  of  any  hard  substances  which  may  pro- 
duce sparks  by  accidental  contact  with  the  tools  used  in  the 
process.  The  rammer  should  not  be  of  metal.  A  stout  hickory 
bar  is  usually  the  tamping-rod.  Frequently  tamping-bags  of 
stout  paper  are  filled  with  the  tamping  material,  when  the  latter 
is  of  low  quality  and  contains  grit.  This  prevents  contact  of  the 
gritty  substance  with  the  rock  surrounding  the  hole.  Anthracite 
miners  use  a  half-inch  iron  bar  clubbed  at  the  end  to  i  J  inches 
for  a  tamper.  This  has  a  groove  at  the  side  to  allow  for  the 
needle. 

For  the  same  reason  the  needle  mentioned  above  must  not  be 
of  iron  or  steel,  but  of  copper.  Sometimes  it  is  merely  a  copper- 
tipped  iron  rod.  In  either  case  it  is  sharply  pointed  and  when 
withdrawn  leaves  a  cavity  for  the  insertion  of  the  fuse.  Black 
powder  is  ignited  by  intense  heat  or  by  a  spark  or  by  de-tonation. 
The  last  named  is  produced  by  the  explosion  of  a  small  quantity  of 
fulminate  of  mercury  in  a  cap  by  conducting  to  it  sufficient  amount 
of  heat  through  a  fuse  ignited  at  the  surface.  The  detonator 
may  be  dispensed  with  and  the  heat  from  the  fuse  alone  may 
be  sufficient  to  ignite  the  powder.  The  result  is  not  so  bene- 
ficial as  when  obtained  by  detonation,  as  shown  by  the  table 
on  page  672.  An  electric  spark  or  detonator  may  also  be 
employed. 

When  the  hole  to  be  fired  is  in  creviced  rock,  which  is  liable 


676  MANUAL  OF  MIX  IXC. 

to  dissipate  the  force  of  the  gases,  a  "bulling-bar"  or  "clay  iron" 
is  used  before  charging  the  hole  to  stuff  soft  clay  into  the  crevices. 

The  Fuse. — The  fuse,  also  called  a  squib,  is  a  thread  of 
powder  wrapped  in  tarred  hemp  or  in  cotton  and  water-proofed 
outside.  When  properly  made  it  burns  uniformly  at  a  rate 
definitely  known  to  the  miner.  One  variety,  for  example,  burns 
at  about  20  inches  per  minute.  A  knowledge  of  the  rate  enables 
the  miner  to  determine  how  short  a  length  of  fuse  may  be  used 
and  still  provide  him  with  safe  escape.  The  fuse  is  supplied 
in  rolls  containing  24  to  40  feet  and  is  cut  off  in  lengths  as  desired. 
It  is  inserted  into  the  cartridge  and  lowered  into  the  hole  with  it 
or  it  may  be  used  in  connection  with  the  needle  as  previously 
described.  It  furnishes  a  means  of  communicating  the  necessary 
spark  and  heat  for  the  ignition  of  the  powder  and  frequently  is 
fitted  with  a  fulminating  cap  at  its  end. 

The  Cap.  —  A  fulminating  cap  is  of  copper  f  inch  long 
and  |  inch  diameter,  having  a  small  quantity  of  fulminate  of 
mercury  deposited  in  it.  This  chemical  explodes  with  heat  with 
violent  fuse  and  ensures  by  its  detonation  the  prompt  ignition 
of  the  powder  in  the  cartridge  into  which  it  is  inserted.  The  cap 
is  slipped  over  the  end  of  the  fuse,  the  top  of  it  being  greased 
with  a  little  cartridge  soap.  Numerous  accidents  have  happened 
from  careless  handling  of  the  caps  which  contain  the  violent 
explosive,  and  care  should  be  taken  that  they  do  not  fall  into 
the  hands  of  the  careless  miner  without  suitable  precaution 
being  exercised.  The  detonation  of  the  explosive  in  the  cap 
increases  the  effect  of  the  black  powder  fourfold  beyond  that 
obtained  by  a  mere  ignition  with  the  use  of  the  simple  fuse,  as 
may  be  noticed  by  the  table  given  above. 

Precautions  in  the  Handling  of  Black  Powder  may  be  sum- 
marized as  follows: 

1.  That  black  powder  and  high  explosives  of  any  kind  are 
not  safe  to  use  together  in  the  same  hole. 

2.  The  cap  is  the  only  exploder  that  is  safe  to  use  in  firing 
off  high  explosives,  which  does  away  entirely  with  tamping. 

3.  The  only  tamping  necessary  on  high  explosives  is  a  small 


BLASTING.  677 

piece  of  miners'  clay,  which  is  easily  put  down  on  the  charge 
with  a  wooden  bar. 

4.  The  practice  of  picking  and  boring  out  missed  holes  that 
have  been  charged  with  either  black  powder  or  high  explosives 
should  be  strictly  prohibited. 

5.  Any  miner  on  the  mine  known  to  use  black  powder  and  high 
explosives  together,  thereby  necessitating  tamping,  will  be  dis- 
charged. 

The  gases  produced  by  the  explosion  of  the  fulminating 
powder  have  the  following  percentages:  CO2,  43;  N,  35;  CO,  12; 
H,  6;  carbohydrates,  4. 

Black  powder  is  employed  for  all  varieties  of  rocks  and  is 
very  serviceable  in  coal.  Its  service  in  the  Galena  veins  and  soft 
rock  is  better  than  obtained  from  the  more  violent  shattering 
explosives,  since  it  does  not  pulverize  the  mineral  so  finely. 

L^wising. — Granite  is  often  quarried  by  a  method  known  as 
le wising.  Several  holes  are  drilled  close  together  and  the  parti- 
tions between  them  broken  down  with  a  flat  steel  bar,  or  broach-bit 
(Fig.  299).  This  extensive  hole  fixes  the  direction  of  the  fracture, 
which  is  usually  selected  as  parallel  to  the  "rift"  or  cleavage. 
Three  drill-holes  make  a  ''complex"  lewis-hole.  The  benefits 
of  this  lewising  may  also  be  secured  by  the  Knox  system,  which 
is  meeting  favor  for  dimension  work.  The  hole  having  been 
drilled,  a  reamer  cuts  V-shaped  grooves  in  its  opposite  sides  to 
determine  the  line  of  break.  The  tamping  is  not  driven  down 
on  the  powder,  but  an  air  space  is  left  between  them.  This 
scheme  permits  expansion  of  the  gases  and  gives  time  to  effect 
iiipture  along  the  plane  desired. 

Powder  Consumption. — The  amount  of  powder  consumed 
for  a  given  work  varies  too  greatly  for  any  general  statement. 
One  foot  of  a  i-inch  hole  is  capable  of  containing  5  ounces  of 
black  powder  or  38  inches  for  a  pound.  The  powder,  to  be 
economically  used,  should  not  fill  more  than  one  third  of  the  hole^ 
In  anthracite  mines  a  keg  of  powder  weighing  25  Ibs.  is  con- 
sumed for  every  40  tons  of  coal.  The  bituminous  miner  breaks 
on  the  average  300  tons  with  a  keg.  The  amount  of  powder 


078  MANUAL  OF  MIX  IXC. 

consumed  per  ton  of  coal  worked  by  hand  and  that  by  machine 
is  about  the  same.  In  longwall  mining  the  consumption  of 
powder  is  very  small. 

Nitroglycerine.  —  Chemically  known  as  trinitroglycerine,  or 
glonoin  oil,  CBH5N3O18.  Made  by  treating  glycerine  to  nitric  and 
sulphuric  acids  at  a  low  temperature.  The  temperature  at  which 
it  will  fire  is  360°.  It  is  exceedingly  sensitive  to  shock  and  will 
explode  under  the  influence  of  a  neighboring  blast.  It  will  not 
take  fire  when  touched  by  a  red-hot  body,  or  if  it  does,  it  burns 
quietly  without  smoke.  If  a  thin  layer  of  it  be  struck  by  a  severe 
blow  it  will  instantly  explode.  If  heated  to  the  temperature  of 
combustion  or  detonated  or  even  jarred  by  a  wave  of  concussion 
its  liquid  mass  will  be  converted  at  once  into  gas  and  if  confined 
will  burst  its  casement.  The  process  of  decomposition  will  be 
complete  and  a  perfect  combustion  of  the  chemical  elements 
will  ensue.  A  small  amount  of  it  explodes  by  detonation,  even 
if  not  buried  in  the  rock,  but  when  lying  on  the  face  of  a  boulder 
in  the  open  air,  the  surrounding  air  will  not  be  able  to  move 
aside,  or  to  take  up  the  rate  of  vibration  of  the  shock  quickly 
enough  to  allow  it  to  expend  its  energy  on  the  air  and  the  rock 
underneath  will  be  badly  shattered. 

The  Manufacture  of  Nitroglycerine. — Several  jars  of  one- 
gallon  size  are  placed  in  a  cool  stream.  Into  them  are  poured 
nitric  acid  and  sulphuric  acid  in  due  proportions.  A  fine  stream 
of  pure  colorless  glycerine  is  then  poured  on  while  stirring  the 
mass.  When  boiling  begins  thick  red  vapors  of  nitrous  gases 
are  emitted.  If  the  glycerine  is  poured  in  too  fast,  or  the  water  is 
not  sufficiently  cool,  the  mixture  will  take  fire  with  a  hissing  noise, 
and  if  it  be  not  stirred  enough  the  blaze  will  shoot  up  several 
feet.  Stirring  is  continued  until  the  action  becomes  less  intense. 
Each  jar  in  turn  receives  a  small  charge  of  glycerine,  after  which 
the  process  is  repeated.  When  no  more  fumes  are  given  off  and 
the  addition  of  glycerine  produces  no  effect,  the  decomposition  is 
complete  and  a  heavy  milky,  oily  fluid  will  be  noticed  on  the 
bottom.  The  acids  are  decanted  and  washed  thoroughly  with 
water.  Into  an  old-fashioned  wooden  churn  the  nitroglycerine 


BLASTING.  679 

is  then  poured  with  an  abundant  supply  of  water  and  churned 
actively  until  litmus  paper  shows  no  acid  reaction.  Upon  the 
freedom  from  all  acids  depends  the  safety  and  the  permanence  of 
the  nitroglycerine.  The  product  being  poured  into  a  wooden 
bucket  with  some  clear  water  its  milkiness  will  soon  disappear  and 
a  colorless  compound  is  the  result.  If  after  a  few  days  it  dis- 
colors the  acid  must  be  removed.  A  yellow  tinge  requires  the 
additional  churning  with  cold  water;  an  orange  color  requires 
some  alkali;  but  if  it  becomes  cloudy  and  deeper  in  tinge  the 
material  is  valueless  and  should  be  either  discarded  or  replaced 
at  once. 

Dynamite.  —  This  explosive  was  discovered  in  1864  by 
: ':'.  Nobel  and  has  received  universal  adoption  in  a  variety  of  forms 
for  almost  all  purposes  of  blasting.  The  term  is  a  generic  one 
and  includes  any  mixture  of  nitroglycerine  with  some  inert  or 
mechanical  absorbent  material.  Earth,  sawdust,  wood  pulp, 
infucorial  earth,  and  magnesia  are  used  as  absorbents,  in  addi- 
tion to  which  are  added  in  some  cases  explosive  bases  which 
increase  the  strength.  Nitroglycerine  is  the  active  principle  of 
the  explosive  and  its  percentage  in  the  compound  determines 
the  power  of  the  explosive.  Infusorial  earth,  which  absorbs  three 
times  its  bulk  of  nitroglycerine,  excels  all  other  bases  and  fur- 
nishes a  very  strong  explosive.  It  is  usually  graded  as  No.  i. 
The  weaker  grades  of  dynamite  are  designated  as  No.  2,  No.  3, 
etc.  The  strength  of  No.  i  is  about  three  quarters  that  of  the 
pure  article  and  six  times  that  of  black  powder.  The  powders 
known  as  Atlas,  dualine,  forcite,  etc.,  differ  only  in  the  nature 
of  the  material  used  for  absorbent. 

The  compound  is  like  moist  brown  sugar  in  color.  It  freezes 
at  46°  F.,  hardening  into  a  white  mass.  It  is  almost  as  sensi- 
tive to  sudden  changes  of  temperature  or  pressure  as  nitrogly- 
cerine. It  is  perfectly  safe  to  handle  and  is  harmless  so  long  as 
it  does  not  exude  from  its  absorbent. 

The  constituents  of  some  of  the  high  explosives  are  indicated 
below: 


680  MANUAL  OF  MINING. 

Tonite  is  macerated  guncotton  52.5  and  baryta  nitrate  47.5. 

Gelatine  is  soluble  guncotton  2.5  and  nitro  97.5. 

Dualine  is  nitro  50,  sawdust  30,  and  nitre  20. 

Rendrock  is  nitro  40,  parafPne  7,  nitre  40,  and  wood  fibre  13. 

Atlas  A  is  nitro  75,  tibreless  wood  21,  nitre  2,  and  magnesia  2. 

Hercules  No.  i  is  nitro  75,  chlorate  of  potash  i,  nitre  2,  sugar  2,  and  magnesia  20. 

Giant  No.  2  is  nitro  40,  rosin  6,  sulphur  6,  absorbent  8,  and  nitre  40. 

Rackarock  is  nitrobenzol  22.3  and  chlorate  potash  77.7. 

Vulcanite  is  mealed  gunpowder  and  nitro  in  different  proportions. 

Experiments  upon  the  relative  efficiency  of  the  various  ex- 
plosives have  been  made  under  water  and  are  recorded  in  General 
Abbot's  "Submarine  Mines,"  but  no  formula?  can  be  prepared 
from  the  results  because  of  the  varying  influences  of  the  foreign 
substances  added  to  the  mail  explosive. 

The  Storage  of  Nitroglycerine.  —  Nitroglycerine  and  its 
various  compounds  may  be  exploded  by  sudden  increase  in 
the  temperature  and  some  by  a  neighboring  shock.  All  are 
liable  to  decompose  under  a  low  heat.  It  will  exude  from  its 
absorbent  when  hot  or  exposed  to  a  wet  atmosphere.  Precau- 
tion must  therefore  be  taken  in  the  storage  and  the  handling  of 
the  material  to  avoid  these  sources  of  danger.  They  should 
be  stored  in  large  cool  dry  caves  or  sheds  well  provided  with 
ventilating  flues  and  remote  from  scenes  of  blasting.  They 
should  not  be  allowed  to  freeze,  for,  while  in  that  state  they 
cannot  be  fired,  thawing  will  be  necessary  before  they  can  be  of 
service.  The  thawing  of  dynamite  introduces  an  element  of 
danger,  unless  care  be  taken  to  avoid  a  sudden  increase  in  tem- 
perature or  an  excess  of  heat.  Dynamite  is  thawed  out  by  slow 
heating  in  a  bucket  immersed  in  another.  In  a  jacket  is  circtK 
latcd  lukewarm  water.  If  the  bucket  be  kept  away  from  the 
direct  fire  or  heat  an  excessive  heating  is  impossible.  Dynamite 
which  has  been  stored  a  long  time  should  be  tested  by  litmus 
paper  for  acid  reaction,  which  indicates  decomposition,  from 
which  spontaneous  combustion  may  ensue.  Less  accidents 
occur  from  the  storage  of  dynamite  or  nitroglycerine  than  from 
the  keeping  of  black  powder. 

Being  a  more  powerful  mixture  than  black  powder,  the  ni'ro- 
glycerine  compounds  are  more  economical  by  reducing  the  labor 


BLASTING.  68 1 

of  the  miner;  requiring  less  tamping;  breaking  the  mineral 
finer;  requiring  smaller  holes;  consuming  less  steel  and  supplies, 
and  materially  shortening'  the  time  of  mining.  As  water  does 
not  injure  it,  it  js  without  rival  as  an  explosive  in  wet  ground, 
and  this  becomes  more  manifest  as  the  rock  encountered  is  harder. 
Methods  of  Charging  with  Nitroglycerine. — The  holes  are 
loaded  with  nitroglycerine  by  pouring  from  a  tin  cup  upon  the 
fuse  with  its  cap  and  covering  the  mass  with  water.  Dynamite 
may  be  loaded  like  black  powder,  the  cartridges  into  which  it  is 
made  being  obtainable  of  any  desired  diameter  and  of  a  length 
of  about  8  inches  each.  As  many  such  cartridges  may  be  inserted 
as  necessary.  Safety-fuses  or  electric  wire  and  cap  are  buried 
in  the  uppermost  cartridge,  which,  after  placing  in  position,  is 
slightly  tamped  with  earth  and  ready  for  ignition. 

Black  powder  is  often  used  in  combination  with  giant  powder 
"to  start  the  hole."  The  proportion  is  usually  one  half  pound 
of  black  powder  to  three  sticks  of  giant.  This  practice  is  often 
questioned,  but  the  fact  that  the  entire  ground  is  broken  to  the 
very  bottom  of  the  hole  without  leaving  any  collar  at  the  surface 
is  sufficient  evidence  of  its  advantage. 

The  Secondary  Explosion  of  Powder. — The  great  disadvantage 
attending  the  use  of  any  form  of  explosive,  particularly  in  coal- 
mines, lies  in  the  production  of  smoke  or  gases  more  or  less 
noxious.  The  combustion  of  carbon  hydrogen  and  sulphur  in 
the  presence  of  the  oxygen  provided  by  the  various  nitrates  added 
to  the  explosive  results  in  the  production  of  carbonic  acid,  car- 
bonic oxide,  sulphurous  acid,  nitrous,  and  other  injurious  gases, 
and  water,  which  necessitate  plentiful  ventilation.  When  the 
combustion  is  perfect  and  complete  all  the  carbon  is  converted 
to  carbonic  acid  and  the  sulphur  to  sulphuric  acid.  But  in 
the  deficiency  of  oxygen  resulting  from  the  addition  of  adul- 
terants or  an  imperfect  mechanical  mixture  of  the  constituents, 
carbonic  oxide  is  developed  as  well  as  smoke,  which,  with  the 
other  injurious  and  combustible  gases,  are  injected  from  the  hole 
into  the  room  or  stope.  Sometimes  this  is  accompanied  with 
a  large  volume  of  sparks.  The  conditions  are  therefore  favorable 


682  MANUAL  OF  MIXIXC. 

for  a  secondary  explosion  in  the  room,  which  might  be  further 
aggravated  by  the  presence  of  coal-dust  or  gas,  and  thus  form  a 
nucleus  for  a  bigger  explosion  after  being  brought  into  contact 
with  neighboring  volumes  of  gas.  For  this  reason  common 
black  powder,  blasting  gelatine,  and  carbonite  are  prohibited 
in  many  coal-mines  inclined  to  be  dusty  or  fiery.  Every  com- 
missioner who  has  examined  into  the  safety  and  efficiency  of 
blasting-powder  has  reported  that  black  powders  are  dangerous 
and  should  be  prohibited.  Some  of  them  are  so  highly  adulter- 
ated that  not  only  is  the  combustion  incomplete,  but  the  gases 
evolved  by  the  combustion  afterwards  dissociate  and  sometimes 
recombine,  forming  dangerous  gases  that  may  incite  secondary 
explosions.  Nitroglycerine  and  dynamite,  which  are  chemical 
compounds  capable  of  complete  and  instantaneous  combustion,  are 
much  safer,  besides  being  more  economical,  since  no  carbonic  oxide 
is  produced.  These  latter  explosives  may  be  employed  to  great 
advantage  and  with  less  risk  even  in  the  dry,  dusty,  and  fiery 
mines. 

Flameless  Explosives. — A  variety  of  explosives  have  been 
made  which  produce  no  carbonic  oxide  and  hence  no  flame, 
and  which  are  known  as  safely  explosives,  sometimes  also  called 
Sprengel  explosives.  These  are  composed  of  a  mixture  either  of 
two  solids,  a  solid  and  a  liquid,  or  two  liquids,  one  of  which 
should  be  a  hydrocarbon  and  the  other  a  compound  rich  in 
oxygen,  neither  at  the  same  time  being  sensitive  to  friction.  Prin- 
cipal among  these  is  a  group  of  explosives — ammonite,  roburite, 
carbodynamite,  and  bellite — in  which  the  chemical  to  supply  the 
oxygen  is  a  nitrate,  as  in  black  powder.  Nitrate  of  ammonia  or 
of  barium  was  most  commonly  employed,  while  the  other  com- 
bustible is  nitronaphthalene  or  dinitrobenzol.  Owing  to  the 
deliquescent  nature  of  ammonium  nitrate  the  two  elements  are 
not  mixed  until  ready  for  immediate  use.  When  so  mixed  the 
explosive  is  dipped  in  melted  wax  to  keep  it  dry. 

A  detonating  wave  such  as  is  produced  by  the  fulmination 
of  mercury  is  alone  capable  of  firing  it.  When  the  two  elerrents 
are  intimately  mixed  in  proper  proportions,  the  shock  of  the 


BLASTING,  683 

detonation  is  communicated  to  the  layers  of  molecules  in  the 
immediate  proximity  of  the  cap  and  the  powder,  whereby  the 
"molecule  edifice"  is  destroyed.  This  initial  force  is  augmented 
to  the  degree  corresponding  to  the  heat  evolved  in  the  decom- 
position until  the  total  is  consumed.  The  decomposition  of  the 
ammonia  nitrate  absorbs  such  an  amount  of  heat  from  that  of 
the  initial  force  as  to  reduce  the  final  temperature  of  the  gases 
developed  and  projected  into  the  air  below  the  temperature  of 
ignition  of  fire-damp.  Owing  to  the  impossibility  of  an  incom- 
plete combustion,  an  excess  of  available  oxygen  is  added  in  an 
excess  of  nitrate  of  ammonia,  which,  absorbing  more  heat,  effectu- 
ally prevents  the  formation  of  carbonic  oxide  and  nitrogen  oxides. 
Changes  of  temperature  do  not  affect  the  mixture,  freezing  for 
three  months  does  not  injure  it,  and  there  is  no  exudation  to 
endanger  it.  Unlike  gunpowder  or  any  nitroglycerine  com- 
pounds, it  will  not  explode  by  percussion,  fire,  or  electric  spark. 
If  struck  with  a  heavy  hammer  the  portion  of  the  explosive 
directly  hit  is  decomposed  by  the  heat  developed  by  the  blow, 
but  the  remainder  is  not  affected.  If  mixed  with  gunpowder  and 
fired  the  latter  explodes,  scattering  the  former  without  affecting  it. 
Detonators. — These  are  required  to  produce  a  violent  con- 
cussion and  an  explosion,  if  otherwise  weak  explosives,  and  are 
of  different  grades,  depending  upon  the  quantity  of  fulminate 
of  mercury  in  them.  The  strongest  contains  23  grains  of  ful- 
minate and  is  known  as  No.  i;  No.  2  contains  15  grains,  and 
No.  3,  8  grains.  The  strength  of  the  detonator  used  increases 
with  the  uncertainties  of  the  two  constituents  of  the  powder. 
Ammonite  containing  between  87  and  89  parts  of  ammonium 
nitrate  and  n  to  13  parts  of  dinitronaphthalene,  and  bellite 
which  contains  79  to  81  parts  of  nitrate  with  19  to  21  'parts  of 
metadinitrobenzol,  require  a  No.  i  detonator.  Roburite  with 
nearly  the  same  amount  of  nitrate,  but  containing  the  more 
sensitive  chloronaphthalene  and  dinitrobenzol,  and  carbonites 
containing  30  to  36  parts  of  barium  nitrate,  25  to  27  parts  of 
nitroglycerine,  and  37  to  43  parts  of  wood  meal,  may  be  fired  by 
a  No.  2  detonator.  Ardeer  powder,  containing  a  little  more 


6S4  MANUAL  OF  MINING. 

nitroglycerine  than  carbonite,  with  n  to  13  parts  of  kieselguhr, 
requires  for  its  explosion  a  No.  3  detonator. 

These  detonators  add  a  new  element  of  danger  to  mining, 
for  in  their  deflagration  a  spluttering  stream  of  sparks  frequently 
results,  which  is  liable  to  ignite  any  coal-dust  which  may  hap- 
pen to  be  present  in  the  room.  This  is  the  main  objection  to 
the  use  of  flameless  explosives,  though  the  latter  are  safer  than 
the  other  forms  of  powder  now  in  use.  Perhaps  the  use  of 
a  detonator  of  still  higher  power  might  reduce  the  danger  by 
producing  total  combustion,  but  it  is  probable  also  that  the  in- 
creased temperature  and  the  heat  evolved  might  be  sufficient  to 
ignite  the  gaseous  mixture  in  the  room. 

Safety-explosives  cost  nearly  twice  that  of  black  powder, 
besides  producing  a  greater  proportion  of  coal-dust. 

Smokeless  Powders. — Powders  which  are  said  to  be  smokeless 
are  made  by  mixing  in  suitable  proportions  guncotton  and 
nitroglycerine,  their  combinations  being  effected  by  the  use  of 
camphor  and  acetones.  Their  energy  is  moderated  by  the  addi- 
tion of  inert  materials  or  by  increasing  the  proportion  of  gun- 
cotton,  though  this  results  in  a  lower  rapidity  of  explosion.  Such 
powders  are  handled  in  small  cubical  grains  and  are  safe.  They 
are  exploded  by  detonation,  not  by  percussion. 

"  Schneiderite "  is  a  powder,  light  yellow  in  color,  quite  oily 
to  the  touch,  forming  lumps  readily  when  pressure  is  applied. 
Considered  alone  "  schneiderite "  is  a  wholly  inert  substance  of 
perfect  stability  and  containing  in  itself  no  explosive  .substance 
whatever.  The  elements  of  which  it  is  composed  only  com- 
bine to  form  an  explosive  at  the  very  moment  of  the  explosion 
under  the  influence  of  a  detonating  primer. 

When  the  detonator  is  not  used  "schneiderite"  may  be  sub- 
mitted to  the  most  violent  shocks  with  impunity.  It  is  not  influ- 
enced by  fire.  Thrust  into  a  fire  it  burns  with  difficulty.  It  is 
also  uninfluenced  by  the  most  extreme  cold.  It  is  sensitive  to 
but  one  alteration  and  that  only  diminishes  its  explosive  qualities; 
this  is  the  alteration  which  may  result  from  its  hygroscopicity. 
To  avoid  the  absorption  of  moisture  it  is  necessary  to  make  sure 


BLASTING.  685 

of  the  imperviousness  of  the  cartridge  papers.  It  is  easy  to 
restore  all  its  properties  by  drying  it  in  a  stove  or  simply  in  the 
sun. 

The  handling  or  the  transportation  of  "  schneiderite "  is  not 
dangerous  under  any  circumstances  or  under  any  conditions  of 
preservation. 

The  Theory  of  Blasting. — When  an  explosive  is  fired,  the 
tension  of  its  gases  acts  uniformly  in  all  directions  upon  its  encase- 
ment. A  tendency  to  rupture  is  developed  along  the  line  offer- 
ing the  least  resistance  to  the  action  of  the  powder.  In  a  rifle 
the  bullet  finds  a  comparatively  easy  line  of  escape  along  the 
barrel.  This  is  the  line  of  least  resistance.  In  rock  the  line  of 
least  resistance  will  be  along  one  of  the  systems  of  cleavage  planes, 
in  the  direction  of  the  shortest  distance  from  the  centre  of  gravity 
to  the  surface  or  along  the  bore-hole  itself. 

It  is  the  purpose  of  the  miner  to  ascertain  the  direction  of 
the  line  of  least  resistance  and  to  drill  the  bore-hole  to  be  loaded 
with  powder  in  a  direction  approximately  at  right  angles  to  it. 
Except  in  massive  rock  presenting  a  plane  surface  for  attack 
this  line  of  fracture  cannot  be  arbitrarily  selected,  for  the  powder 
will  seek  its  easiest  vent  for  escape  from  the  explosive  chamber. 
For  this  reason  a  hole  cannot  be  drilled  along  the  line  of  least 
resistance.  It  becomes  merely  a  cannon;  the  powder  blows  out 
without  other  result. 

The  Line  cf  Least  Resistance. — In  Drinker's  "Explosive 
Compound"  will  be  found  a  table  showing  the  relative  resistances 
of  different  rocks  and  coefficients  representing  their  toughness. 
By  its  use  may  be  determined  the  quantity  of  a  given  grade  of 
powder  necessary  to  produce  the  rupturing  effect.  Let  W  be 
the  weight  in  ounces  of  the  blasting-agent,  L  the  distance  in 
feet  of  its  centre  of  gravity  from  the  surface  of  the  working-face, 
and  C  the  coefficient  of  the  rock  resistance,  then  W=CL3.  In 
a  given  mine  the  value  of  C  may  be  determined  by  a  series  of 
trials,  the  other  two  variables  being  altered  in  amount,  employing 
as  little  explosive  and  placing  the  hole  for  as  long  a  line  L  as 
possible.  Thus  if  a  27-inch  hole  breaking  rock  along  a  line  of 


686 


MANUAL  OF  MINING. 


20  inches  length  shows  an  average  consumption  of  5  ounces  dyna- 
mite, then  C  becomes  1.09.  A  subsequent  hole  with  a  line  of 
least  resistance  of  30  inches  would  require  17  ounces  of  the  same 
explosive. 

The  line  of  least  resistance,  which  is  the  line  of  general  throw, 
extends  from  a  point  approximately  the  centre  of  the  charge 
to  the    nearest  external  point,  from  a  to  b  in  Figs.  322  to  325. 
FIG.  322.  FIG.  323.  FIG.  324.  FIG.  325. 

• 


Illustrating  the  Line  of  Least  Resistance. 

In  soft  rock  \vhose  cohesion  is  comparatively  slight,  the  line  of 
least  resistance  may  be  a  long  one — nearly  that  of  the  depth  of 
the  hole  (Fig.  322).  In  tough  rock  the  line  ab  is  short  (Fig.  323). 
The  volume  of  rock  broken  down  is  proportional  to  the  weight 
of  explosive  used  and  to  the  cube  of  the  line  of  least  resistance. 
With  a  given  depth  of  hole  and  of  charge  in  it  the  length  of  line 
is  fixed.  The"  volume  of  rock  broken  is  determined.  To  increase 
the  effectiveness  of  the  powder  the  line  must  be  increased  by 
varying  the  direction  of  holes  or  the  amount  of  powder  diminished. 
A  deep  hole  loaded  with  a  powerful  explosive  is  Jience  more 
economical  than  shallow  holes  or  low  r^ade  of  Dowder. 


. 

FIG.  326. — Shooting  in  Tight  Ground. 


When  the  line  of  least  resistance  exceeds  in  length  that  of 
the  hole,  as  in  Fig.  326,  no  execution  is  done,  other  than  that 


BLASTING.  687 

indicated  by  the  result  at  K.  The  inclination  of  the  drill-hole 
with  a  flush  working-face  cannot  ordinarily  exceed  45°  when  black 
powder  is  used.  It  may  be  deflected  to  60°  if  the  rock  is  soft 
and  the  explosive  powerful. 

Blasting  in  Creviced  Rocks. — The  cleavage  planes  of  the 
stratified  rocks  furnish  numerous  lines  of  least  resistance  which 
may  be  advantageously  employed  as  lines  of  rupture.  Wedges 
or  picks  are  often  inserted  to  split  FIG 
them.  Certain  coals  break  freely 
along  these  cleavage  planes.  In 
some  of  the  massive  rocks  the 
rifts  and  seams  are  sufficiently 
pronounced  to  afford  lines  of  weak 

Drilling  in  Creviced  Rock. 

resistance  to  rupture.     The  miner 

avails  himself  of  these  faults  in  the  rock  and  drills  the  hole  normal 
to  them  but  not  down  to  the  seams  in  Fig.  327.  If  the  centre 
of  gravity  of  the  powder  is  well  below  the  crevice  b,  the  block 
above  a  will  be  loosened  for  a  considerable  distance  on  either  side 
of  the  hole. 

In  the  case  illustrated  in  Fig.  328  the  line  of  least  resistance 
is  across  to  the  free  face  of  rock,  but  there  is  likewise  a  tendency 
to  split  along  the  plane  b ;  the  rock  will  doubtless  break  downward 
to  the  right.  In  Fig.  261  the  shot  m  has  broken  to  the  bedding 
plane,  and  somewhat  shattered  the  rock  ahead.  The  shot  n 
and  that  at  m  (Fig.  262)  will  break  to  the  black  clay  gouge.  That 
at  o  will  probably  blow  out  without  any  execution.  The  shot  p 
will  blow  out  a  large  block  if  those  near  o  have  been  properly 
fired.  By  firing  the  holes  in  order  from  above  downward  greater 
extension  is  effected,  as  the  bedding  planes  assist  the  miner.  The 
shot  g  in  Fig.  263  is  well  placed  to  break  the  rock  at  the  plane 
of  cleavage.  With  the  stratification  dipping  down  from'  the 
face  the  shots  proceed  in  order  upward. 

In  coal  and  soft  or  brittle  metalliferous  minerals  a  weak  explo- 
sive is  used  to  avoid  injury  to  the  product.  When  a  thin  vein 
of  the  latter  adjoins  a  hard  rock  the  usual  practice  is  to  blow 
down  the  rock  with  lightly  loaded  holes.  In  shafts  or  tunnels 


688  MANUAL  OF  MINING. 

no  attention  is  paid  to  the  condition  in  which  the  mineral  is  broken 
down,  as  rapid  progress  is  sought.  For  simultaneous  firing  no 
heed  is  paid  to  the  economy  of  the  line  of  least  resistance.  Such 
holes  are  machine-drilled  and  are  limited  in  inclination. 

Blasting  in  Massive  Rocks. — In  homogeneous  massive  rocks 
there  is  no  pronounced  crevice  or  seam  to  assist  the  miner  in  the 
displacement  of  large  masses.     The  line  of  least  resistance  must 
be  created  by  properly  directing  the  drill  to 
L  blow  down  the  maximum  possible  volume  of 

rock  with  the  minimum  of  drilling.     But  an 
additional  problem  is  presented.     The  miner 
must  consider  the  manner  in  which  working- 
m%        face  will  be  left  after  the  shot.     Each  shot  in 
turn  becomes  a  "bearing-in"  shot  for  its  suc- 
cessor (Fig.  329).    Without  any  planes  of  cleav- 
FIG.  329.— Bearing-     age  to  which  to  break  porphyry,  vein  quartz, 
and   granites,  the  direction    of    the    shooting 
must  be  to  their  working-face.     This  problem  is  of  special  diffi- 
culty when  the  working-face  is  the  small  area  of  a  tunnel  or  shaft. 
Such  an  explosive  is  known  in  miners'  parlance  as  "  tight  ground." 
Considerable  skill  is  necessary  to  accomplish  much  work  under 
such  circumstances.     The  usual  practice  consists  in  so  breaking 
the  ground  as  to  avoid  plane  surfaces,  but  produce  irregularities  of 
prone  such  as  are  illustrated  in  Figs.  324  and  330  to  332.     Rupture 
FIG.  330.  FIG.  331.  FIG.  332. 


Taking  Advantage  of  Irregularities  of  Face. 

may  then  be  effected  along  the  dotted  lines  toward  the  recesses 
in  the  rock.     Judgment  must  be  exercised  in  directing  the  holes; 


BLASTING. 


689 


as,  for  example,  hole  2  in  Fig.  301  will  displace  more  rock  than 
No.  i  with  equal  powder  and  work.  Indeed  it  may  happen 
that  the  powder  will  simply  pulverize  the  rock  along  the  line  of 
break  of  No.  i,  dislodging  the  bowlder  alone.  So  in  Fig.  260 
is  illustrated  a  poorly  directed  hole.  If  drilled  to  k  the  blast 
will  take  place  along  ks,  but  only  a  small  charge  is  needed  for  the 
short  line.  As  drilling  is  an  expensive  operation  it  would 
have  been  more  economical  to  have  drilled  it  less  vertically. 
Had  the  operations  stopped  near  m  no  more  powder  would  be 
required  to  blow  out  at  bm  than  at  sk,  though  difficulty  would 
have  been  experienced  later  in  blasting  out  bmks. 

Expanding-bits. — A  soft  or  brittle  rock  may  be  blown  out  in 
large  masses  by  excavating  a  large  chamber  at  the  bottom  of  a 
FIG.  333.  FIG.  334. 


An  Expanding-bit. 

very  deep  hole  and  rilling  it  with  black  powder,  sufficient  for  the 
work.      The  chamber  may  be  produced  by  an  expanding-bit 


690 


MANUAL  OF  MINING. 


(Fig.  333)  or  by  a  small  charge  of  powder.  When  the  desired 
depth  of  hole  has  been  reached  a  pair  of  cutter- wings  are  forced 
out,  which  when  revolved  cut  out  a  chamber.  Powder  accom- 
plishes the  same  end  by  shattering  the  rock.  This  method  is 
known  as  squibbing. 

Simultaneous  Firing. — It  has  been  shown  that  increased 
effectiveness  is  obtained  from  an  explosive  by  the  use  of  per- 
fect detonators.  A  still  further  gain  results  from  the  ignition 
of  the  shots  simultaneously,  by  which  they  materially  assist  one 
another  in  effecting  disruption.  The  concussion  produced  by 
the  explosive,  though  progressing  radially  in  every  direction 
from  the  centre  of  the  charge,  succeeds  in  rupturing  the  rock 
only  over  a  limited  volume,  as  is  represented  in  Fig.  335. 


0 

b 

/ 

IF 

x          .  *' 

'\              >X 

-^t£— 

T 

i 

i 

1 
1  . 

FIG.  335- 

Either  shot  (i  or  /)  drilled  into  the  block  would  break  out 
only  a  triangular  prism.  The  hole  i  would  break  the  prism 
bounded  by  the  parallel  faces  CAd  and  eif,  or  the  hole  ;  would 
break  the  prism  bounded  by  the  faces  ABD  and  IFJ.  The 
rock  is  broken  toward  the  free  face,  but  the  force  expanding 
laterally  and  backward  into  the  solid  ground  is  entirely  lost, 
the  amount  lost  increasing  with  the  intensity  of  the  explosion. 
If,  however,  several  contiguous  holes  be  so  placed  that  their  lateral 
rupturing  tendencies  will  superpose,  they  will  shatter  rock  which 
by  the  single  shooting  would  only  have  been  fractured.  The 
holes  i  and  j  would  break  the  mass  of  rock  included  between 
the  trapezoids  ABGC  and  IJHE.  The  latter  mass  is  mani- 
festly much  in  excess  of  the  aggregate  product  of  the  single  holes. 


BLASTING. 


691 


The  zone  of  an  explosive  effect  may  be  divided  into  three  sec- 
tions— a  sphere  of  pulverization  immediately  surrounding  the 
bottom  of  the  hole,  a  cone  of  rupture  with  its  apex  near  the 
centre  of  gravity  of  the  cartridge  and  its  base  of  an  area  dependent 
upon  the  relative  explosive  and  resisting  powers,  and  a  zone  of 
vibrations  extending  into  the  rock  where  the  undulations  are  too 
weak  to  produce  more  than  an  occasional  fracture.  By  simul- 
taneous firing,  the  explosives  may  be  placed  at  such  distances 
apart  as  to  permit  the  overlapping  of  the  zones  of  vibration,, 
whose  cumulative  effect  may  produce  rupture. 

The  effect  of  synchronous  firing  is  about  1.4  times  as  great 
as  that  obtained  for  the  same  amount  of  powder  fired  in  consecu- 
tive shots.  Benjamin  Frost  says  in  his  report  on  the  Hoosac 
Tunnel  that  "greater  depths  of  holes  are  admissible"  and  greater 
advance  obtained. 


The  sole  means  of  obtaining  perfect  simultaneity  is  by  elec- 
tricity; fuses  will  not  burn  with  sufficient  regularity  to  be  relied 
upon.  The  holes  are  charged,  the  cap  is  inserted,  two  wires- 


692 


MANUAL  OF  MINING. 


take  the  place  of  fuse,  and  the  tamping  is  done.  The  wires  are 
separated,  each  one  being  connected  with  its  neighbor  in  such 
manner  (Fig.  336)  that  a  continuous-wire  circuit  E  extends 
throughout  the  series  of  holes.  A  is  the  fuse,  B  the  cartridge, 
and  C  the  cap. 

The  Electric  Fuse. — The  several  holes  to  be  fired  simultane- 
ously are  charged  with  an  electric  fuse,  A  (Fig.  337),  filled  v.ith 
fulminates,  B\  the  latter  is  detonated  by  the 
incandescence  of  a  fine  platinum  thread,  E, 
from  which  two  copper  wires,  C,  lead.  The 
fuse-wires  are  of  standard  lengths,  usually  4  or 
6  feet  for  hand-driven  holes  and  16  or  18  feet 
for  machine- drilled  holes. 

The  ends  of  the  fuse-wires  are  cleaned  and 
twisted  on  the  connecting  wires  D  (Fig.  336), 
and  when  all  these  have  been  connected,  the 
circuit  is  completed  by  the  final  connections  with 
the  machine.  The  leading  wire  E  is  usually 
wound  on  reels  to  pay  out  a  supply  as  needed 
for  the  work.  The  connecting  .wires  are  of  com- 
mon copper  wire  and  no  larger  in  diameter  than 
is  necessary  to  transmit  the  current  freely.  By 
a  judicious  division  of  the  mine  into  districts  the 
amount  of  wiring  may  be  materially  reduced. 

In  Fig.  339  is  shown  a  section  of  the  magneto 
machine,  which  consists  of  a  weak  principal 
magnet,  A,  with  an  armature,  B,  revolving 
between  its  poles;  said  armature  is  rotated  by 
the  rack-and-pinion,  C,  operated  from  the 
handle  on  the  down  stroke.  As  the  rack-bar 
strikes  the  spring,  D,  at  the  bottom,  it  breaks  the  continuous  cur- 
rent between  the  two  platinum  bearings,  E,  and  causes  it  to  pass 
through  the  outside  circuit,  the  fuses,  etc.  A  machine  which  will 
fire  16  shots  at  one  time  costs  $25;  one  good  for  70  can  be  bought 
for  $55. 

Much  wire  is  buried  in  the  debris  of  shooting.     In  a  mine 


FIG.  337-— The 
Electric  Fuse. 


BLASTING. 


'693 


firing  regularly  80  shots  at  a  time,  the  loss  from  this  source  is 
2  Ibs.  per  day;  in  a  small  double- track  mine  tunnel,  2  ounces 
of  connecting  wire  and  7  exploders  per  lineal  foot;  and  in  a 
large  railroad  tunnel,  0.32  feet  of  connecting  wire,  0.07  feet 
leading  wire,  and  0.7  exploder  per  cubic  yard  of  rock  removed. 
FIG.  338.  FIG.  339. 


A  Magneto  Machine  and  its  Mechanism. 

When  the  fuses  have  been  connected,  the  last  man  to  retire 
raises  the  handle  (Fig.  338)  and  with  one  plunge  ignites  afl 
the  shots.  An  increased  efficiency  of  the  explosive  is  not  the 
sole  advantage  of  this  method  of  firing.  It  eliminates  the  numer- 
ous accidents  due  to  premature  explosions  or  to  delayed  shots* 
There  is  but  one  electric  current  and  one  igniting  source.  Fail- 
ing to  fire  them,  there  can  be  no  ignition  later.  It  saves  the 
time  otherwise  wasted  by  the  men,  who  with  their  neighbors  must 
retire  periodically  to  shelter.  It  permits  prompt  firing  from  a 


694 


MANUAL  OF  3//AV.VC. 


FIG.  340. — The  Cyclone  Powder  Loader. 


BLASTING.  695 

safe  distance.    It  is  equally  adapted  to  black  powder  and  to 
dynamite  and  to  flameless  powders. 

REFERENCES. 

Experiments  with  Safety  Explosives,  Bergassessor  Winkhaus,  Trans. 
M.&M.  Eng.,  Vol.  XL VI,  17;  Safety  Explosives,  Bergassessor  Winkhaus, 
Trans.  M.  &M.,  Vol.  XLV  (2),  141;  Flameless  Explosives,  Report  of  Com- 
mittee, A.  C.  Kayll,  Coll.  Eng.,  April  1896,  208;  List  of  Flameless  Explo- 
sives, Copy  of  Orders,  Coll.  Guard.,  June  1897,  1129;  Report  on  Flameless 
Explosives,  A.  C.  Kayll,  E.  &  M.  Jour.,  Vol.  LVIII,  556. 

Blasting  in  Fiery  Mines,  Franz  Brzezowski,  Coll.  Eng.,  April  1896,  209; 
The  Effects  of  Different  Explosives  on  Coal  Dust,  Winkhaus,  Coll.  Eng., 
1896,  39;  Explosives  for  Coal  Mines,  Vivian  B.  Lewis,  Coll.  Eng.,  Vol.  XVI, 
150;  Water  Cartridges  in  Blasting,  L.  Jarolijmek,  Coll.  Guard.,  Vol.  LXXI, 
162;  Elements  of  Defectiveness  in  Shot -firing,  reprint,  Coll.  Guard.,  Dec. 
1894,  1088;  Diminished  Use  of  Explosives  in  Belgian  Collieries,  Victor 
Watteyne,  Coll.  Guard.,  Aug.  1895,  299;  Use  and  Value  of  Explosives, 
C.  J.  Thompson,  Coll.  Mgr.,  Jan.  1896,  15;  Testing  Explosives  for  Coal 
Mines,  E.  &  M.  Jour.,  Vol.  LXI,  567. 

Modern  Development  of  Explosives,  V.  B.  Lewis,  Coll.  Eng.,  March 
1895;  Nature  and  Use  of  Industrial  Explosives,  J.  Daniell,  Coll.  Guard., 
March  15,  1895,  497. 

Electric  Ignition  of  Blasting  Explosives,  Coll.  Guard.,  Oct.  1896,  792; 
Electric  Ignition  of  Blasting  Explosives,  B.  Heise,  Coll.  Guard.,  March 
189?)  598>  Electric  Firing,  J.  von  Lauer,  Coll.  Guard.,  Jan.  1897,  161;  Shot- 
firing  by  Electricity,  P.  Mehers,  Coll.  Guard.,  1894,  241. 

Dangers  of  Percussion  Fuses,  Trans.  M.  &  M.  Eng.,  Vol.  XLVI,  41; 
Percussion  Fuses  and  Their  Suitability  in  Fiery  Mines,  J.  von  Lauer,  Coll. 
Guard.,  1896,  258. 

Detonators,  Coll.  Eng.,  Dec.  1895;  Explosives  and  Detonators,  J.  House, 
Coll.  Guard.,  March  22,  1895,  546. 

Relative  Cost  and  Efficiency  of  Powder,  Coll.  Eng.,  May  1897,  459; 
Influence  of  Diameter  of  Holes  in  Blasting,  J.  Daniell,  Coll.  Guard.,  Dec. 
1894,  1088;  Experiments  for  Ascertaining  the  Comparative  Effect  of  Ex- 
plosives, Bergassessor  Winkhaus,  Coll.  Guard.,  Sept.  1895,  539- 

Dynamite  Accidents  and  Prevention,  James  Ashworth,  Coll.  Guard.,  Nov. 
1896,  927;  High  Explosives,  W.  J-  Orsman,  Coll.  Guard.,  Jan.  1897,  168; 
Explosives,  W.  T-  Orsman,  Coll.  Guard.,  Nov.  1804,  835;  Explosives  in  Bel- 
gium, Victor  Wattevne,  Coll.  Gu~rd.,  1807,  208;  Prevention  of  Accidents 
from  High  Exolosives,  James  Ashworth,  Coll.  Guard.,  Nov.  1896,  927. 

Cost    of   Mining   Items,    Calif.  Bureau  Mineralogy,    loth,   852;    Coal- 


696  MANUAL   OF  MINING. 

mine  Accounts,  Min.  Industry,  Vol.  IV,  207;  Mine  Accounting,  Walter  M. 
Jeffrey,  M.  &  M.,  Oct.  1903;  The  Colliery  Cost  Sheets,  J.  J.  Priest,  Fed. 
Inst.  M.  E.,  Vol.  IX;  Mining  Problems  in  Colorado,  Percy  A.  Leonard, 
Min.  Sci.  &  Pr.,  Jan.  30,  1904;  Mining  Errors,  F.  Danvers,  Iron  and  Coal 
Trades  Rev.,  March  30,  1900;  The  Training  of  a  Mining  Engineer,  R.  A. 
S.  Redmayne,  Coll.  Guard.,  Dec.  19,  1903;  A  Practical  System  of  Mine 
Accounting,  E.  Jacobs,  Eng.  Mag.,  May  1903. 

Regarding  the  Output  of  the  Mine,  M.  &  M.,  Vol.  XXIII,  51;  Ore  in 
Sight,  M.  &  M.,  Vol.  XXIII,  13;  The  Leasing  System,  Prof.  Arthur  Lakes, 
M.  &  M.,  Nov.  1903. 

Why  Some  Mines  are  Abandoned,  M.  &  M.,  Vol.  XXIV,  565;  Loss 
Reduction  in  the  Life  of  a  Mine,  M.  &  M.,  Vol.  XXIII,  66. 

What  Constitutes  a  Mine,  M.  &  M.,  Vol.  XXIII,  319;  Mining  Engi- 
neering and  Valuation  of  Mines,  G.  W.  Miller,  Min.  Sci.  Pr.,  April  n, 
1903;  Daily  Examination  of  Mines,  James  Freer,  111.  Min.  Inst.,  Vol.  I,  183; 
Notes  on  Mine  Sampling  on  the  Main  Reef  Series,  D.  JT  Williams,  Jour. 
Chem.  Met.  Soc.  of  S.  Africa,  Oct.  1902;  The  Valuation  of  Iron  Ores  and 
other  Raw  Materials,  Bernard  Osann,  Stahl  u.  Eisen,  Oct.  i  and  15,  1902. 

Anthracite  Mine  Surveying  Practice,  Eng.  &  Min.  Jour.,  Feb.  n,  1904. 

Colliery  Costs,  Coll.  Guard.,  Vol.  LXXIX,  1028;  Pumping  by  Com- 
pressed Air,  Coll.  Guard.,  Vol.  LXXXI,  571. 

Explosives,  Henry  Louis,  Coll.  Guard.,  Feb.  1897,  302;  Use  and  Abuse 
of  Dynamite,  A.  Slaght,  Bureau  of  Mines,  Ontario,  1895,  285;  Manufacture, 
Use,  and  Abuse  of  Dynamite,  Harry  A.  Lee,  E.  &  M.  Jour.,  LXI,  182. 


GLOSSARY   OF  MINING  TERMS. 

Absolute  pressure.    The  pressure  reckoned  from  a  vacuum. 

Absolute  temperature.  The  temperature  reckoned  from  —461° 
F.  or  -273°C. 

Adit.  A  horizontal  passage  from  daylight  into  a  mine,  and 
en  or  along  the  vein.  It  differs  from  a  tunnel  in  that  it  is  not 
through  country  rock  across  to  the  vein;  it  differs  from  a,  drift, 
which  has  neither  end  in  daylight. 

Aerophone.  A  respirator  in  the  form  of  a  tank,  receiving 
the  exhalations  from  the  lungs,  which  contains  chemicals  designed 
to  revive  the  air  and  render  it  fit  for  breathing.  It  is  used  by 
rescuers  after  mine  accidents. 

After-damp.  An  irrespirable  gas  remaining  after  an  explo- 
sion of  fire-damp. 

Anemometer.  An  appliance  for  measuring  the  velocity  of 
an  air-current. 

Anticlinal.     A  fold  of  the  rock  or  strata,  convex  upward. 

Apex.    The  edge  of  a  vein  nearest  to  the  surface. 

Back  pressure.  The  loss,  expressed  in  pounds  per  square 
inch,  due  to  getting  the  steam  out  of  the  cylinder  after  it  has  done 
its  work. 

Balance-bob.  A  heavy  triangular  truss,  the  long  horizontal 
arm  of  which  supports  a  weight  at  one  end  ballasted  to  balance 
the  weight  of  pump-rods  at  the  other. 

Barrier  pillars.  Pillars,  larger  than  the  ordinary,  left  at 
stated  intervals  to  prevent  the  crushing  of  the  roof  from  extend- 
ing beyond  the  section  enclosed  by  them. 

Bar-timbering.  A  system  of  supporting  a  tunnel  roof  by 

697 


6gB  MANUAL  OF  MINING. 

long  top  bars  while  the  whole  lower  tunnel- core  is  taken  out, 
leaving  an  open  space  for  the  masons  to  run  up  the  arching. 
Under  certain  conditions  the  bars  are  withdrawn  after  the  masonry 
is  completed,  otherwise  they  are  bricked  in  and  not  drawn. 

Battery.  The  lower  platform  of  a  coal-chute.  A  term  also- 
used  to  define  a  timber  bulkhead  in  a  gallery. 

Bearing-up  stop.  A  partition  of  brattice  or  plank  that  serves 
to  conduct  air  to  a  face. 

Bed.    A  seam  of  mineral  occurring  among  the  stratified  rock, 

Bench.  The  divisions  of  a  coal-bed  caused  by  seams  of  clay 
or  slate;  also  used  to  express  the  artificial  divisions  in  the  process 
of  mining;  also  a  terrace  at  the  outcrop  of  a  seam. 

Black-damp.     Carbonic  acid  gas. 

Blast.  The  operation  of  forcing  air  by  blowing.  The  opera- 
tion of  exploding  powder  or  other  agents. 

Blind  drijt.  A  horizontal  passage  in  the  mine  not  yet  con- 
nected with  the  other  workings. 

Blind  lead  or  blind  lode.     A  vein  having  no  outcrop. 

Block  coal.     Coal  that  breaks  freely  into  rectangular  blocks. 

Blossom  rock.  The  rock  detached  from  a  vein,  but  which 
has  not  been  transported. 

Blower.  A  discharge  of  gas  from  coal;  also  a  fan  for  forcing 
air  into  a  mine. 

Blow-out.    The  decomposed  mineral  exposure  of  a  vein. 

Bluff.     Blunt. 

Bob.     See  Balance-bob. 

Brattice.  A  canvas  or  plank  partition,  nailed  to  posts  longi- 
tudinally, with  a  level  or  shaft  to  divide  the  same  into  two  com- 
partments for  the  purposes  of  separating  two  air-currents. 

Break-through.  A  narrow  passage  cut  through  a  pillar  con- 
necting rooms. 

Breast.    The  face  of  a  gallery  or  heading 

Buggy.  A  small  mine-wagon  for  conveying  coal  from  face 
to  gangway. 

Bulling-bar.  An  iron  bar  used  to  pound  clay  into  the  crevices 
crossing  a  bore-hole,  which  is  thus  rendered  gas-tight. 


GLOSSARY   OF  MINING  TERMS.  699 

Bull-pump.  A  single-acting  direct  pump  consisting  of  a 
steam-cylinder  placed  over  the  shaft.  The  steam  drives  its 
piston,  to  which  the  pump-rods  are  attached.  By  means  of  the 
piston  attached  to  the  rods  the  water  is  lifted  by  the  steam  pressure. 
The  down  stroke  is  effected  by  the  weight  of  the  pump-rods. 

Bunions.  Timbers  placed  horizontally  across  a  shaft.  They 
serve  to  brace  the  wall-plates  of  the  shaft-lining,  and  also,  by 
means  of  planks  nailed  to  them,  to  form  separate  compartments 
for  hoisting  or  ladder  ways. 

Butt.     The  end  faces  of  coal. 

Butt-entry.  The  gallery  driven  at  right  angles  with  the 
butt-joint. 

Butty.     A  miner  working  on  contract. 

Cage.     An  elevator  platform. 

Cap.  A  term  signifying  the  point  at  which  a  vein  is  con- 
tracted; it  is  also  the  rock  covering  the  ore. 

Cartridge.     A  paper  tube  filled  with  an  explosive. 

Casing.     The  tubing  of  a  well-hole  to  prevent  caving. 

Centre.  A  temporary  support,  serving  at  the  same  time  as- 
a  guide  to  the  masons,  placed  under  an  arch  during  the  progress 
of  its  construction. 

Chain  pillar.  An  untouched  block  of  mineral  on  either  side 
of  the  gangway. 

Choke-damp.     Carbonic  acid  gas. 

Chute.  An  inclined  trough  or  timbered  shaft  through  which 
ore  is  delivered  by  gravity  to  a  receptacle  below.  See  also  Shoot. 

Clack.     A  pump-valve. 

Clearance.  The  space  between  the  piston  at  the  end  of  its 
stroke  and  the  valve  face,  or  the  end  of  the  cylinder. 

Cleat.  A  joint  produced  by  the  natural  tendency  of  coal 
or  rock  to  cleave  or  split  in  a  certain  direction  not  parallel  to 
the  plane  of  bedding. 

Coal-measures.     The  strata  embracing  the  carboniferous  coals. 

Column- pipe.  The  line  of  pipe  through  which  the  mine- 
water  is  pumped. 

Compression.     In  steam  practice,  the  action  of  the  piston  in 


7°0  MANUAL  CF  MINING. 

compressing  the  steam  remaining  in  the  cylinder,  after  the  closure 
of  exhaust-valves,  into  the  clearance  space. 

Contact  vein.  A  vein  lying  between  two  dissimilar  rock 
masses  or  strata. 

Counter.    A  cross  vein. 

Counterbalance;  Counterpoise.  A  weight  used  to  balance 
another  weight  or  the  vibrating  parts  of  machinery. 

Counter  gangway.  One  which  is  driven  diagonally  to  the 
rise  until  the  workings  are  reached,  when  it  turns  off  parallel 
to  the  main  haulageway. 

Country  rock.  The  main  rock  of  the  region  through  which 
the  veins  cut,  or  that  surrounding  the  veins. 

Crab.     An  iron  windlass  for  moving  heavy  weights. 

Creep.  The  crushing  of  the  overlying  rock  resulting  in  the 
floor  rising. 

Crib.  A  framework  built  like  a  log-cabin.  It  may  be  a 
mere  pillar  afterwards  filled  with  rock,  or  it  may  be  the  lining 
to  a  mill-hole  or  shaft. 

Cross-course.    An  intersecting  vein. 

Cross-cut.  A  horizontal  passage  driven  across  the  country 
rock  to  a  vein. 

Cross-heading.  A  transverse  drift  which  is  driven  for  pur- 
poses of  ventilation  from  one  gangway  to  another. 

Culm.  The  fine  waste  of  coal-mines  containing  dirt  as  well 
as  coal-dust. 

Curb.  A  timber  frame  intended  as  a  support  or  foundation 
for  the  lining  of  a  shaft. 

Cutter.  A  term  employed  in  speaking  of  any  coal-  or  rock- 
cutting  machines,  the  men  operating  them,  or  the  men  engaged 
in  underholing  by  pick  or  drill. 

Dam.     A  bulkhead. 

Day-shift.  The  gang  of  miners  working  during  the  day- 
time. • 

Dead.  The  valueless  matter  of  a  vein,  also  gangue,  or  waste. 
It  is  usually  made  the  stowing  or  filling  of  an  excavated  por- 
tion of  the  seam,  bed,  or  vein,  and  then  is  called  gob,  waste,  or 


GLOSSARY  OF  MINING   TERMS.  701 

stull  dirt.  Sometimes  the  term  is  employed  in  speaking  of  a 
sluggish  ventilating  current. 

Dead  quartz.     Quartz  carrying  no  mineral. 

Dead  work.  Exploratory  or  prospecting  work  that  is  not 
directly  productive. 

Deposit.     Irregular  ore  bodies — not  veins. 

Dip.  The  angle,  measured  by  the  steepest  line  in  the  plane 
of  a  layer  of  rock,  from  the  horizon. 

Dirt- fault.  A  partial  replacement  of  coal  in  a  seam  by  clay. 
Not  a  true  fault. 

Drag.  The  point  of  union  of  two  veins  which  meet  with- 
out intersecting. 

Drift.    Any  subterranean  horizontal  passage.     See  Adit. 

Driving.     Excavating  drifts,  adits,  or  levels. 

Drum.    The  cylinder  on  which  a  hoisting-cable  is  wound. 

Dumb  drift.  A  gallery  which  conducts  the  air  around  a 
ventilating  furnace  to  the  up-cast  shaft. 

Dump.  The  pile  of  rock  which  has  been  hoisted  to  the  sur- 
face and  deposited  there.  It  may  be  said  to  be  a  low-grade 
ore  reserve. 

Duty.  The  unit  of  measure  of  the  work  of  a  pumping-engine 
expressed  in  foot-pounds  of  work  obtained  from  a  bushel  or 
100  Ibs.  of  fuel. 

Dyke.    A  fissure  filled  with  igneous  matter. 

Empty  or  empties.  An  unloaded  car  or  the  track  along  which 
it  travels. 

Exploder.  A  chemical  employed  for  the  instantaneous  ex- 
plosion of  powder. 

Exploitation.  The  working  of  a  mine  and  similar  under- 
takings; the  examination  instituted  for  that  purpose. 

Eye.  The  hole  in  a  pick-  or  hammer-head  for  receiving  the 
handle. 

Face.  The  exposure  of  rock  at  which  work  is  being  per- 
formed. 

Fathom.  A  volume  of  rock  equal  to  six  feet  square  multi- 
plied by  the  thickness  of  the  vein. 


7° 2  MANUAL  OF  MINING. 

Fault.  The  dislocation  of  a  vein.  The  term  is  also  im- 
properly used  in  coal-seams.  See  also  Rock-fault  or  Dirt-fault. 

Fire-damp.  A  carburetted  hydrogen  gas,  inflammable  and 
specifically  lighter  than  air. 

Fire-setting.  The  process  of  exposing  very  hard  rock  to 
intense  heat,  rendering  it  thereby  easier  of  breaking  down. 

Fissure  vein.  Any  mineralized  crevice  in  the  rock  of  very 
great  depth. 

Float.  Broken  and  transported  particles  or  bowlders  of  vein, 
matter. 

Floor.    The  stratum  below  a  mineral  bed. 

Foot-wall.    The  face  of  rock  below  the  vein. 

Forepoling.    A  mode  of  timbering. 

Free.    A  term  employed  in  speaking  of  loose  mineral. 

Fuse.  A  tube,  ribbon,  or  wire  filled  or  saturated  with  a 
combustible  compound,  used  for  exploding  powder. 

Gad.    An  iron  or  steel  wedge;   a  chisel-bit  pick. 

Gallery.    A  horizontal  passage. 

Gallows- frame.  The  frame  supporting  a  pulley,  over  which 
the  hoisting- rope  passes  to  the  engine. 

Gangue.    The  barren  portion  of  the  vein. 

Gangway.    The  principal  level  of  a  coal-mine. 

Gash-vein.  A  mineralized  fissure  that  extends  only  a  short 
distance  vertically.  It  may  be  confined  to  a  single  stratum  of 
rock,  but  is  a  comparatively  shallow  vein. 

Gateway.    A  gangway  having  ventilating  doors. 

Gauge  pressure.  The  pressure  shown  by  an  ordinary  steam- 
gauge.  It  is  the  absolute  pressure  plus  that  of  the  atmosphere. 

Goaf.  The  excavated  space  of  a  coal-mine,  usually  filled 
with  the  valueless  portion  (gob}  of  the  seam. 

Gobbing-up.     Filling  with  waste. 

Gouge.  The  layer  of  clay  or  decomposed  rock  which  lies 
along  the  wall  or  walls  of  a  vein.  It  is  not  always  valueless. 

Guide.  The  timbers  nailed  to  the  timbers  of  a  shaft  for  the 
purpose  of  guiding  the  cage. 

Gunboat.    A  skip;  a  self-dumping  box  used  in  slopes. 


GLOSSARY  OF  MINING   TERMS.  703 

Hanging-wall.    The  wall  of  rock  above  the  vein. 

Head-gear.     A  derrick. 

Heading.  A  drift  or  airway.  The  section  of  tunnel  driven 
in  advance  of  the  lower  section  or  bench. 

Heave.    A  dislocation  of  the  strata. 

Helve.    A  handle. 

Hewer.     A  coal- miner. 

Hitch.  A  dislocation  of  a  vein.  Also,  a  shoulder  or  hollow 
cut  in  the  rock  to  support  one  end  of  a  stull  or  other  timber. 

Hogback.     See  Horse. 

Holing.  The  picking  of  a  groove  in  the  lower  part  of  a  coal- 
seam  for  the  purpose  of  facilitating  the  breaking  down  of  the 
upper  mass. 

Horse.  A  mass  of  country  rock  lying  within  a  vein.  Any 
irregularity  cutting  out  a  portion  of  the  vein.  See  Dirt-fault 
and  Rock- fault. 

H-piece.  The  portion  of  a  column  pipe  containing  the 
valves  of  the  pump. 

Incline.  An  entry  into  a  mine  following  the  dip  of  the  vein 
or  seam. 

Indicator.     That  which  points  or  directs. 

Some  forms  show  the  position  of  the  cave  in  the  shaft.  Others 
record  upon  paper  the  pressure  of  the  steam  in  an  engine-cylinder 
.at  various  points  in  the  piston-stroke. 

Inplace.     A  vein  or  deposit  in  its  original  position. 

Intake.  The  entry  which  conducts  the  incoming  air-current 
to  the  mine.  It  is  synonymous  with  downcast. 

Jar.  That  part  of  the  drilling  apparatus  which  takes  up 
the  shock  of  impact  of  the  falling  tools  upon  the  bottom  of  the 
hole. 

Kibble.     An  iron  ore-bucket. 

Kirving.     See  Holing. 

Lagging.  The  slabs  or  small  timber  placed  between  the 
main  timber  sets  and  the  roof  or  walls  to  prevent  small  rock 
from  falling  into  the  drift. 

Lath.     A  plank  laid  over  a  framed  centre  or  used  in  poling. 


7 °4  MANUAL  OF  MINING. 

Laundry-box.  The  box  at  the  surface  receiving  the  water 
pumped  up  from  below. 

Level.  A  horizontal  passage  in  a  vein-mine,  numbered  i, 
2,  3,  etc.,  consecutively  from  the  surface,  or  from  the  cross-cut 
tunnel,  down. 

Lift.  All  the  mine  workings  connected  with,  opened  from, 
and  mined  out  at  one  level;  also  the  length  of  pump-pipe  between 
stations. 

Live  quartz.  A  variety  of  quartz  usually  associated  with 
mineral. 

Location.     A  mining  claim. 

Lode.    A  mineralized  fissure. 

Longwall.  The  system  of  mining  coal  without  leaving  any 
pillars. 

Man-engine.  A  mechanical  appliance  for  raising  and  lower- 
ing miners. 

Manhole.  A  hole  or  an  auxiliary  shaft  through  which  a 
man  may  pass  in  going  from  one  level  to  another,  into  a  stope,. 
or  from  one  ladder  to  another. 

Measures.  A  term  embracing  the  strata  of  a  geological 
series. 

Mill-hole.  An  auxiliary  shaft  connecting  a  stope  or  other 
excavation  with  the  level  below. 

Mill-run.  A  test  of  the  value  of  a  quantity  of  ore  as  distin- 
guished from  an  assay,  which  tests  "pocket  specimens." 

Mineral.  Any  constituent  of  the  earth's  crust  that  has  a 
definite  composition. 

Mineral  oil.  Petroleum  or  other  liquid  obtained  from  the 
earth. 

Miner's  inch.  The  unit  of  measurement  of  water  used  by 
the  sluice-miners.  It  is  that  amount  of  water  hourly  discharged 
through  an  opening  i  inch  square  under  a  head  of  several  inches. 
If  the  head  is  7  inches  and  the  hole  is  through  a  plank  2  inches 
thick,  a  miner's  inch  is  equal  to  about  90  cubic  feet  per  hour. 

Mining.  In  its  broad  sense  embraces  all  that  is  concerned 
with  the  production  of  minerals  and  their  complete  utilization. 


GLOSSARY  OF  MINING   TERMS.  705 

Mining  retreating.  A  process  of  mining  by  which  the  vein 
is  untouched  until  after  all  the  gangways,  etc.,  are  driven,  when 
the  mineral  extraction  begins  at  the  boundary  and  progresses 
toward  the  shaft. 

Moil.  A  short  length  of  steel  rod  tapered  to  a  point,  used 
for  cutting  hitches,  etc. 

Narrow  work.  Working  places  narrower  than  the  rooms, 
entries,  headings,  break-throughs,  gangways,  etc. 

Nitro.  A  corrupted  abbreviation  for  nitroglycerine  or  dyna- 
mite. 

Ore.  A  mineral  of  sufficient  value  (as  to  quality  and  quan- 
tity) which  may  be  mined  with  profit. 

Ore-shoot.  A  large  and  usually  rich  aggregation  of  mineral 
in  a  vein.  Distinguished  from  pay-streak  in  that  it  is  a 
more  or  less  vertical  zone  or  chimney  of  rich  vein  matter 
extending  from  wall  to  wall  and  having  a  definite  width 
laterally. 

Outcrop.    The  exposed  portion  of  a  vein  on  the  surface. 

Outlet.     An  exit  passage  from  the  mine. 

Output.     The  product  of  a  mine. 

Panel.  The  division  of  a  mine  which  is  isolated  from  neigh- 
boring districts  and  provided  with  distinctive  haulage  and  mining 
systems. 

Parting.  A  joint  in  the  rock,  or  a  crevice  in  a  seam,  filled 
with  clay  or  slate ;  a  switch  or  turnout  to  allow  loaded  and  empty 
cars  to  pass  one  another. 

Pay-streak.  The  thin  layer  of  a  vein  which  contains  the 
pay-ore. 

Pike.     A  pick. 

Pinch.    A  contraction  in  the  vein. 

Pipe.  An  elongated  body  of  mineral.  Also  the  name  given 
to  the  fossil  trunks  of  trees  found  in  coal-veins. 

Pit.     A  shaft. 

Pitch.     See  Dip. 

Pit-man.  The  shaft-man  who  attends  to  the  shaft  equip- 
ments, pumps,  etc/ 


706  MANUAL  OF  MINING. 

Placer.  A  surface  accumulation  of  mineral  in  the  wash  of 
streams. 

Plane.  An  inclined  tramway  for  lowering  cars  by  gravity  or 
raising  them  by  means  of  a  stationary  engine. 

Plat.  A  platform.  A  swinging  or  revolving  door  used  in- 
termittently to  connect  two  trackways. 

Plumb  (adjective).     Vertical. 

Plummet.  A  string  or  fine  copper  wire  attached  to  a  heavy 
weight;  used  for  determining  the  vertically  of  shaft-timbering. 

Plunger.    The  solid  piston  of  a  force-pump. 

Pocket.     A  rich  and  large  body  of  ore  in  the  vein. 

Poling.  The  process  of  timbering  by  the  use  of  poles,  for 
timbering  in  soft  ground. 

Poll-pick.    A  combination  pick-  and  hammer-head. 

Poppet;  also  puppet.  A  pulley-frame  or  the  head-gear 
over  a  shaft.  A  valve  which  lifts  bodily  from  its  seat  instead 
of  being  hinged. 

Post  and  stall.     See  Pillar  and  Room. 

Power-drill.  A  rock-drill  employing  steam,  air,  or  electricity 
as  a  motor. 

Prerelease.  The  act  of  discharging  steam  or  air  from  the 
cylinder  before  the  piston  has  reached  the  end  of  its  stroke. 

Prop.  A  piece  of  timber  or  metal  placed  normally  to  the 
roof  or  wall  for  its  support. 

Prospect.  The  name  given  to  underground  workings  the 
value  of  which  has  not  yet  been  made  manifest.  A  prospect 
is  to  a  mine  what  mineral  is  to  ore. 

Prospecting.  The  process  of  seeking  pay-ore  or  the  prelim- 
inary operations  of  a  mine. 

Pump.     Any  mechanism  for  raising  water  out  of  a  mine. 

Quick  (adjective).  Soft,  running  ground;  an  ore  or  pay- 
streak  is  said  to  be  quickening  when  the  associated  minerals 
indicate  richer  mineral  ahead. 

Rafter  timbering.  That  in  which  the  timbers  appear  like 
roof-rafters. 

Rgacher.    A  slim  prop  reaching  from  one  wall  to  the  other. 


GLOSSARY   OF  MINING   TERMS.  707 

Reamer.    An  enlarging  tool. 

Reef.  The  outcrop  of  a  hard  vein  projecting  above  the  sur- 
face. Also  applied  to  auriferous  quartz  lodes. 

Regulator.  A  sliding- door  to  apportion  the  amount  of  air 
to  be  admitted  into  a  section  of  the  mine. 

Rib.     A  pillar  of  vein-matter  left  to  support  roof  or  wall. 

Robbing.     The  taking  of  mineral  from  pillars. 

Rock- fault.  A  disturbed  portion  of  a  vein  in  which  coal  is 
replaced  by  sandstone. 

Roof.     The  stratum  overhead. 

Room.  A  working  place  in  a  flat  mine;  corresponds  to  slope 
in  a  steep  vein. 

Run.  A  mode  of  contract  work  in  which  steep  parts  of  coal- 
seams  are  driven  and  paid  for  by  the  lineal  foot  or  yard  of  progress. 

Saddle.     The  ridge,  of  a  stratum  or  ore-bed. 

Safety-cage.     One  supplied  with  safety  appliances. 

Safety-lamp.  A  lamp  in  which  the  flame  is  protected  from 
immediate  contact  with  the  surrounding  atmosphere. 

Salting.  Placing  foreign  ore  in  the  crevices  of  a  vein  or  else- 
where to  fraudulently  raise  its  apparent  value. 

Samson-post.  An  upright  supporting  the  working-beam 
which  communicates  oscillatory  motion  to  pump  or  drill-rod. 

Sand- pump.     See  Sludger. 

Scale.  The  incrustation  deposited  in  boilers  from  evapo- 
rated waters. 

Scraper.     A  tool  for  cleaning  out  drill-holes. 

Seam.     A  layer  of  mineral. 

Seed-bag.  A  water-tight  packing  of  flaxseed  around  the 
tube  of  a  drill-hole  to  prevent  the  influx  into  the  hole  of  water 
from  above. 

Selvage.     See  Gouge. 

Sett  or  Set.     A  frame  of  timber. 

Shaft.     A  vertical  opening  from  the  surface. 

Shaly.     Brittle  ground. 

Sheave.    A  grooved  wheel  over  which  a  rope  is  turned. 

Shell- pump.     See  Sludger. 


708  MANUAL  OF  MINING. 

Shelly.     Broken  ground. 

Shijt.     The  duration  of  day's  work — from  six  to  ten  hours. 

Shoot.     To  break  rock  by  means  of  explosives. 

Shute.     An  inclined  boardway  through  which  coal  is  delivered. 

Sill.  The  floor-piece  of  a  timber  sett,  or  that  on  which  the 
track  rests. 

Slack.     Small  dirt  or  coal. 

Slate.     Bony  coal  and  hard  clay. 

Slickenside.      The  polished  surface  of  the  vein  or  its  walls. 

Slitter.    See  Pick. 

Slope.  An  incline.  It  is  an  inside  slope  when  it  does  not 
extend  to  the  surface. 

Sludger.  A  cylinder  having  an  upward  opening  valve  at 
the  bottom,  which  is  lowered  into  a  bore-hole  to  pump  out  the 
sludge  or  fine  rock  resulting  from  drillings. 

Smift.    A  slow-burning  fuse. 

Snore.  The  hole  in  the  lower  part  of  a  sinking  or  Cornish 
pump  through  which  water  enters. 

Sollar.  The  plank  flooring  of  a  gallery  covering  a  gutter- 
way  beneath.  Also  the  platform  in  a  shaft  between  two  ladders. 

Spears.     Pump-rods. 

Spilling.     A  process  of  timbering  through  soft  ground. 

Spoon.  A  slender  iron  rod  with  a  cup- shaped  projection  at 
right  angles  to  the  rod,  used  for  scraping  drillings  out  of  a  bore- 
hole. 

Sprag.  A  billet  of  wood  used  to  block  the  wheels  of  a  car 
and  check  its  speed.  Sprags  are  permanently  used  on  self- 
acting  inclines.  A  very  steep  line  requires  a  sprag  in  each  of 
the  four  wheels,  while  on  a  moderate  pitch  only  one  may  be 
necessary  to  block  a  hind  wheel.  Also,  a  short  prop. 

Square-sett.    A  variety  of  timbering  for  large  excavations. 

Squeeze.  The  closing  of  a  room  by  the  settling  of  the  roof 
or  the  rising  of  the  floor.  The  thinning  away  of  a  seam. 

Squib.     A  slow  fuse  used  for  igniting  an  explosive. 

Stall.    The  longwall  working- face,  or  a  room. 

Station.     An  excavation  adjoining  a  shaft  for  receiving  the 


GLOSSARY  OF  MINING   TERMS.  709 

pump  or  V  balance-bob,  or  for  landing  the  hoisting  convey- 
ances. 

Stockwerke.  A  mass  of  country  rock  so  impregnated  by  a 
congeries  of  veins  as  that  the  whole  must  be  mined  together. 

Slope.  A  step.  The  excavation  of  a  vein  in  a  series  of 
steps. 

Sloping  overhand.  Mining  a  stope  upward,  the  flight  of 
steps  being  reversed. 

Sloping  underhand.  Mining  a  stope  downward  in  such  a 
series  that  presents  the  appearance  of  a  flight  of  steps. 

Slowing.  The  debris  of  a  vein  thrown  back  of  a  miner  to 
support  the  roof  or  hanging  wall  of  an  excavation. 

Strike.  The  bearing  of  a  horizontal  line  through  the  middle 
of  a  vein. 

String-rods.  A  line  of  surface  rods  connected  rigidly  for 
the  transmission  of  power;  used  for  operating  small  pumps  in 
adjoining  shafts  from  a  central  station. 

Slull.  A  stick  of  timber  or  platform  for  supporting  miners 
or  vein-waste  temporarily  or  permanently. 

Slull  dirt  or  slull  rock.     Material  supported  upon  the  stulls. 

Stump.  A  pillar  between  the  gangway  and  its  parallel  air- 
way. 

Slylhe.     Carbonic  acid  gas. 

Sump.  The  lowest  point  of  the  workings,  usually  a  pro- 
longation of  a  shaft,  into  which  the  mine  water  is  drained  and 
from  which  it  is  pumped. 

Swamp.     A  trough-shaped  basin  in  a  coal-mine. 

Synclinal.    The  depression  of  a  seam  or  stratum. 

Tail  rope.  The  secondary  rope  used  for  balance,  which  is 
attached  underneath  the  cages  of  a  hoisting-plant,  or  at  the  tail 
end  of  the  loaded  and  empty  trains  &f  cars  on  a  slope  for  rais- 
ing the  empty  cars  or  skips. 

Tamping.  The  process  of  making  a  bore-hole  gas-tight  by 
the  use  of  clay. 

Tempering.  The  act  of  reheating  and  properly  cooling  a 
bar  of  metal  to  any  desired  degree  of  elasticity. 


710  MANUAL  OF  MINING. 

Through*  or  Thirling.  A  passage  cut  through  a  pillar  to 
connect  two  rooms. 

Throw.     The  amount  of  dislocation  of  a  vein. 

Tram.     The  pair  of  parallel  lines  of  rails  of  a  trackway. 

Trammer.     One  who  pushes  cars  along  the  track. 

Trap.  A  door  used  for  cutting  off  a  ventilating  current, 
which  is  occasionally  opened  for  haulage  or  passage;  guarded 
by  a  trapper. 

Trend.    The  course  of  a  vein. 

Tribute.  A  system  of  contract  mining  by  which  the  miner 
receives  his  pay  out  of  the  gross  value  of  the  ore  sold,  less  a 
certain  deduction  for  royalty  to  the  mine-owner. 

Trolley.  A  small  carriage  truck  having  no  body.  A  trav- 
eller making  connection  between  two  electric  wires. 

Tubbing.     An  iron  o    wooden  cylindrical  lining  of  shafts. 

Tubing.     The  tube-lining  of  bore-holes. 

Tunnel.  A  horizontal  passage;  properly  speaking,  one  with 
both  ends  open  to  the  surface,  but  is  applied  to  one  opening 
at  daylight  and  extending  across  the  country  rock  to  the  vein 
or  mine. 

Underhand  work.     Picking  or  drilling  downward. 

Underholing.     See  Holing  and  Kirving. 

Upcast.    An  entry  through  which  the  air-current  rises. 

Upraise.  An  auxiliary  shaft,  a  mill-hole,  carried  from  one 
level  up  toward  another. 

Vein.     A  mineral  deposit  filling  a  fissure  or  crevice. 

Vug.     A  cavity  in  the  rock. 

Wagon  breast.  One  from  which  ore  or  coal  can  be  carried 
by  wagon. 

Wall.     The  faces  of  a  fissure;   the  sides  of  a  gallery. 

Wall-plate.  The  long  horizontal  stick  in  a  shaft  timbering 
frame  which  is  parallel  to  the  vein. 

Waste.    The  debris  of  an  excavation;  gob;  goaf. 

Wedging.  The  material,  moss  or  wood,  used  to  render  the 
shaft-lining  tight. 

Whim.    A   hoisting  appliance  consisting  of  pulley  support- 


GLOSSARY  OF  MINING   TERMS.  711 

ing  the  hoisting-rope  which  is  wound  on  a  drum  turned  by  a 
beam  attached  to  a  horse. 

Whip.  A  hoisting  appliance  consisting  of  a  pulley  support- 
ing the  hoisting-rope  to  which  the  horse  is  directly  attached. 

White-damp.     The  noxious  gas  called  carbonic  oxide  gas. 

Winch  or  Windlass.  A  hoisting-machine  consisting  of  a 
horizontal  drum  operated  by  crank-arm  and  manual  labor. 

Winning.     Recovering  or  mining. 

Winze.  A  small  auxiliary  underground  shaft  sunk  from  an 
upper  level. 

Wire-drawing.  The  operation,  accidental  or  otherwise,  of 
reducing  the  pressure  of  steam  between  the  boiler  and  the  cyl- 
inder. 

Working-barrel.  The  cylinder  in  which  the  pump-piston 
operates. 

Workings.  Any  underground  development  from  which  ore 
is  being  extracted. 


SIGNALLING. 

The  writer  would  suggest  the  following  code  of  signalling: 
The  engineer  should  signal  below  when  he  is  ready  to  hoist  by 
raising  the  bucket  or  cage  a  foot  or  two  and  lowering  again. 
This  is  important,  particularly  for  the  safety  of  the  man  who 
may  be  engaged  in  igniting  the  blasts  near  to  the  place  of  hoist. 
The  strokes  should  be  made  at  regular  intervals  and  a  halt  of 
a  few  seconds  between  the  signals.  Thus 

5  bells-4  bells  would  mean  send  tools  to  the  fourth  level. 

1  bell:   To  hoist  or  to  stop  if  in  motion. 

2  bells:  To  lower. 

2-1-1  bells:  No  more  hoisting. 

2-2-2  bells:  To  change  buckets  from  ore  to  water. 

3  bells:    Man  on  board;    lower  or  hoist  slow. 

4  bells:    Stop  or  start  the  pump. 

2-2  bells:    Stop  or  start  the  air-compressor. 

5  bells:   Send  tools  down. 

6  bells:  Send  timbers  down. 

USEFUL  INFORMATION. 

The  area  of  a  circle  is  0.7854  (diameter)2. 

Ratio  of  area  to  circumference  is  as  its  radius  is  to  2. 

An  acre  is  43,560  square  feet. 

A  ton  contains  2000  Ibs.  or  29,166$  troy  ounces. 

A  troy  pound  =  0.82 285  7  avoirdupois  pound. 

A  troy  ounce  =  43 7. 5  grains. 

An  avoirdupois  pound  =  7000  grains. 

A  troy  pound  =  5  760  grains. 

A  cubic  foot  of  gold  =  $300,000. 

A  cubic  foot  of  silver  =  $10,000. 

A  long  ton  is  2240  Ibs.;   a  short  ton  2000  Ibs. 

712 


USEFUL  INFORMATION.  713 

A  bushel  of  coke =40  Ibs. 

A  bushel  of  bituminous  coal  =  76  Ibs.  =  2688  cu.  in. 

1000  feet  (board  measure)  of  dry  white  pine  =  4000  Ibs. 

1000  feet  (board  measure)  of  green  white  pine  =  6000  Ibs. 

One  cord  of  seasoned  wood  =128  cu.  ft. 

A  mile  of  track  (rails  16  Ibs.  per  yd.)  weighs  25  tons  320  Ibs. 

A  mile  of  track  (rails  25  Ibs.  per  yd.)  weighs  39  tons  640  Ibs. 

A  mile  of  track  (rails  35  Ibs.  per  yd.)  weighs  55  tons  o  Ibs. 

A  mile  of  track  requires  9  kegs  (1780  Ibs.)  of  3-in.  spikes. 

A  mile  of  track  requires  15  kegs  (3110  Ibs.)  of  4^-in.  spikes. 

A  mile  of  track  requires  20  kegs  (3960  Ibs.)  of  4^-in.  spikes. 

A  mile  of  track  requires  2640  cross-ties  (2  ft.  apart). 

A  mile  of  track  requires  528  splice-joints  (2  bars,  4  bolts  and 
nuts  per  joint),  each  weighing  5  to  10  Ibs. 

One  miner's  inch  =  21 59  cu.  ft.  per  24  hours  =  0.025  cu-  &• 
per  sec. 

The  pressure  of  i  atmosphere  is  represented  by  a  column 
of  mercury  at  32°  F.,  of  a  height  equal  to  0.760  metre  =  29.92 
inches  =  14. 701  Ibs.  per  sq.  in.,  or  by  a  column  of  water  33.94 
ft.  high. 

i  Ib.  air  at    32°  F.,  14.7  Ibs.  pressure,  occupies  12.387  cu.  ft. 

i   "      "       62°  F.,    "      "         "  "        13.141      " 

i  "      "      i8o°F.,    "      "         "  "        16.106      " 

i  "      "      212°  F.,    "      "         "  "        16.910      " 

i   "water at 62° F.,     "      "          "  "          0.016      " 

i  cu.  ft.  water  at  39°  F.  =  7.49  gallons,  weighs  62.425  Ibs. 

i      "          "      "   62°F.=    "         "  "      62.355    " 

i  gallon      "     "   62°F.  =  23i  cu.  in.,         "        8.325    " 

i  cu.  in.     "     "   62°  F.=  0.00434  gal.,       "        0.036    " 

i  ton          "      "  62°  F.  =  240.3  gals.,  occupies  35.9  cu.  ft. 

A  col.  of  water,     i  sq.  in.  base,  33.947  ft.  high  =  i  atmos. 

"     "    "      "          i       "         "     27.7      in.  high  =  i  Ib.  press. 

"    "    "      "         i       "        "     i  ft.  high          =  0.434  Ib.  pr. 

"     "    "  mercury,  i       "         "     i  in.    "  =0.4914    " 

A  bar  wro't  iron,  i       "     area,  i  ft.  long  =3-33  Ibs.  wt. 

"    "        "       "      i  in.  diam.,       i  "      "  =2.618      " 


MANUAL  OF  MINING. 


TABLE   OF  WEIGHTS   OF  VARIOUS   SUBSTANCES. 


Weight 
per 
Cu.  Ft. 

Cu.  Ft. 
per 
Long 
Ton. 

Weight  per 
Cu.  Ft 

Cu.  Ft.  per 
Long  Ton. 

Gold 

I   87 

Syenite 

Lead  

7IO 

1.  1C 

Porphyry  

1  66  to  171 

13.5  to  13   i 

Silver  

62  c 

3-42 

Slate  

162  to  178 

13.8  to  12  6 

Rolled  iron  

480 

4.68 

Quartz  

16?  .  2 

i-?  6 

Galena  

468 

4.82 

Sandstone  . 

130  to  157 

Nickel  glance.  .  .  . 

468 

4.82 
5  60 

Brick  
Clav 

125  to  135 

18.1  to  16 
18  7 

Chalcocite  
Magnetite  
Specular  iron  ore. 

355-7 
338.6 
327-4 

6.30 
6.63 
6.84 

Anthracite  
Bituminous  
Cannel  

85.4  to  99 
75  to  83 

75 

26.2  tO  22.6 

29.8  to  26.1 
29.8 

Pyrites  
Barytes  
Chalcopyrite  

312 
277-5 

262      I 

7-os 
8.07 
8  cc 

Lignite  
Oak  

Ash 

78  to  84 
73 

28.7  to  27 

30.6 
42  7 

Zinc  blende  

2  CQ 

896 

White  pine 

CQ    C. 

8  06 

^8   7 

57  8 

Limestone  

168 

13-3 

Wood  charcoal.  . 

25  to  39 

89.61057.4 

EQUIVALENTS   OF  FRENCH  AND   ENGLISH  MEASURES. 

ABBREVIATIONS. — M.  =  metre;     cm.  =  centimetre;     G.  =  gramme;     L.  =  litre; 
ft.  =  foot;    lb.  =  pound;    in.  =  inch;    oz.  =  ounce;    dwt.  =  penny  weight;    gr.  =  grain; 


yd.  =  yard;  gal.  =  gallon;  T.  =  troy;    A.  =  avoirdupois;   sec.  =  second;    sq.  =  square; 
cu.  =  cubic;  h.-u.  =  heat-unit. 

M. 

=  3.28  ft. 

atmosphere 

=  0.760  M. 

M. 

=  39.39  in. 

ft.  per  sec. 

=  0.305  M.  per  sec. 

ft. 

=  0.3048  M. 

mile  per  hour 

=  0.447  M-  per  sec. 

in. 

=  0.0254  M. 

M.  per  sec. 

=  3.281  ft.  per  sec. 

yd. 

=  0.9144  M. 

ooo  M.  per  hour 

=  0.621  mile  per  hour. 

Gunter's  chain  =  20.1168  M. 

ooo  G.  per  sq.  M.  =  0.205  Ib.  per  sq.  ft. 

mile 

=  1609.35  M- 

ooo  G.  per  cu.  M.  =  0.0624  Ib.  per  cu.  ft. 

sq.  M. 

=  1.2  sq.  yd. 

Ib.  per  sq.  ft. 

=  4883  G.  per  sq.  M. 

sq.  yd. 

=  0.836  sq.  M. 

Ib.  per  sq.  in. 

=  703077.0  G.  per  sq.  M. 

sq.  in. 
sq.  M. 

=  0.00065  S{1-  M. 
=  1555.2  sq.  in. 

ton  per  ft. 
gal.  per  sq.  ft. 

=  3,333.333  G.  per  M. 
=  48.905  L.  per  sq.  M. 

acre 

=  4048  sq.  M. 

L.  per  sq.  M. 

=  0.0204  gal.  per  sq.  ft. 

cu.  in. 

=  0.0000164  cu-  M. 

G.  per  L. 

=  70.116  gr.  per  gal. 

cu.  ft. 

=  0.02832  cu.  M. 

Ib.  per  cu.  ft.  • 

=  16020  G.  per  cu.  M. 

cu.  M. 

=  1.31  cu.  yd. 

cu.  ft.  per  Ib. 

=  0.0624  cu.  M.  per  1000  G. 

G. 

=  15-43  gr- 

degree  Fahr. 

=  0.5555  deg-  centigrade. 

G. 

=  0.0022  Ib.  A. 

degree  Cent. 

=  1.8  deg.  Fahrenheit. 

T.gr. 

=  0.0648  G. 

Ib.  per  sq.  ft. 

=  column  of  mercury  0.00359 

T.  Ib. 

=  576ogr. 

M.  high. 

T.  Ib. 

=  373.242  G. 

L.  of  normal  air 

=  19.955  grains. 

A.  Ib. 

=  453-593  G. 

G.  M. 

=  0.007233  ft.-lb. 

looo  G. 

=  2  Ib.,  8  oz.,  3  dwt., 

ft.-lb. 

=  138.2  G.  M. 

0.35  gr.  T. 

72  ft.-lbs. 

=  106700  G.  M. 

1000  G. 

=  2  Ib.,  3  oz.,  4  dr., 

calorie 

=  3.968  heat-units. 

10.473  gr-  A. 

heat-unit 

=  0.252  calorie. 

i  fluid  oz. 

=  0.02957  L. 

thermal  unit 

=  0.4536  calorie. 

i  quart 
i  gal. 

=  0.9464  L. 
=  3.78543  L. 

h.-u.  per  Ib.         =0.5555  calorie  per  1000  G. 
calorie  per  1000  G.  =  1.8  h.-u.  per  Ib. 

INDEX. 


ABSOLUTE  pressure,  308 

temperature,  308 
Access  to  workings,  19,  69,  249 
Accidents,  causes,  483,  682 

car,  492 

drilling,  492 

statistics,  488 
Adiabatic  curve,  309 
Adit,  22,  577 

Adjustable  doors,  229,  463 
Aerial  tramways,  299 
Aerophores,  501 
Affidavit  of  labor,  15 
After-damp,  394 
Age  of  veins,  6 
Air-bridges,  45,  459 
Air,  compressed,  307 

compression  of,  307 

-current,  455 

-drills,  579,  637 

expansion  of,  333 

friction  of,  329,  441,  448 

-pumps,  369 

-receivers,  328 

-supply,  96,  447 

-valves,  326 

velocity,  456 

-ways,  407,  442 

weight,  308.  388,  411,  713 
Alignment  of  shafts,  22,  510 

of  tunnels,  578 

Alternating  electric  currents,  172 
American  system  of  tunelling,  583, 

650 

Ammonite.  682 
Ampere,  67 

Anderson  system  of  tunelling,  594 
Anemometers,  455 
Animal  haulage,  268 
Ankylostomiasis,  495 
Anthracite  mining  waste,  63 
Apex,  14,  1 8 


Arches,  in  drifts,  565 

in  tunnels,  392 
Armature,  170 
Asphyxiation,  395 
Assessment,  15 
Atlas  powder,  680 
Atmosphere,  composition,  388 
Atmospheric  pressure,  309,  411,  713 

tension,  308,  411 
Attle,  see  Waste. 
Auger,  stem,  599 
Austrian  method  of  tunnelling,  584 

BAILING-TANK,  338 
Balance-bob,  289,  352 
Balancing  the  resistance,  448 
Barometerrelationofexplosions,396 
Barrier  pillar,  447 
Bar  timbering,  587 
Battery  dam,  56,  567 

for  blasting,  691 
Beard-Mackie  lamps,  399,  474 
Bearing  in,  42,  620,  655,  686 
Bedded  vein,  4,  7,  37 
Behr's  dumping  device,  248 
Belgian  lamp,  476 

tunnel  system,  5:84 
Bellite,  436 

Bending  ropes,  152,  201,  291 
Bit,  convex  diamond,  607 

of  percussion  drill,  614,  625,  644 

of  rotary  drill,  619,  625 
Bituminous  coal-mining, 3 7, 387, 658 
Black  damp,  384 
Black  powder,  674 
Blacksmithing,  629 
Blanket  vein,  17-20 
Blasting,  670 

agents,  496 

off  the  solid,  620 

with  electricity,  277,  435,  691 

with  lime,  671 

715 


7i6 


INDEX. 


Bleichert  tramway,  304 
Blockholing,  624 
Blocking  out  the  mine,  24,  70,  76 
Block  system  of  mining,  586 
Blowers  of  gas,  100 
Blowing- fans,  100,  391,  421 
Bob,  348 
Boiler,  88 

evaporation,  89 

scale,  92 

sectional,  99 
Bonneted  lamps,  476 
Bore-holes,  597 

for  prospecting,  597,  613 
Boring  methods,  537 
Brain's  system  of  drilling,  653 
Brattice,  457 

Break-through,  47,  57,  465 
Breasts,  56,  64,  70 

drill,  618,  661 
Brine  evaporation,  29 
Broach-bit,  636 
Brown's  panel  system,  61 
Buggy  roads,  56 
Bull  pump,  347 

wheel,  599 
Buntons,  513 
Burleigh  air-compressors,  326 

drills,  642 
Butt  headings,  47,  59 

CABLES,  289 

Cage,  safety,  213,  226 

Calculating  depth  of  engine  service, 

hauling  capacity,  158,  276,    280, 
282 

power  transmission  by  air,  212, 
256,  312,  315,  328,  334 

pump  capacity,  373 

size  of  engine,  158 

ventilating  power,  412,  429 

work  of  compressing  air,  312 
Calorific  of  value  fuel,  96 
Calumet  and  Hecla  mine,  498 
Calix  drill,  614 
Cameron  pump,  207 
Camphausen  system,  149 
Candles,  468 
Caps,  fulminating,  676 

in  lamp,  398 

timber,  552 
Carbonic  acid,  388 
Carbonic  oxide,  389,  395 
Carbonite,  682 
Carriage,  225 
Cars,  231,  238,  242 
Car- wheels,  153 
Cataract  engine,  348 


Caving  system,  65,  79 
Cementing  shafts,  513,  535 
Centre-cut  system  of  drilling,  509, 

650 

Centre  props,  552 
Centrifugal  pump,  32,  377 

ventilators,  429 
Chain  pillars,  58 
Chairs,  230 
Channelers,  635 
Chapman  drill,  615 
Chimney  draft,  100,  413 
Chimney,  Evasee,  426 
Chocks,  550 
Chute,  55,  78,  82 
Circular  mils,  169 
Claim,  mining,  14-16 
Clanny  lamp,  455,  472 
Clearance,  113,  327 
Cleat,  27,  407,  520,  687 
Cleavage,  27,  52 
Clifford  lamp,  473 
Clip-pulleys,  288,  302 
Closed  running  fans,  425 
Clutches,  friction,  138,  228 
Coal-beds,  38 

-cutters,  655 

-dust,  493 
Coal  in  lead  veins,  8 

mined  per  fatality,  306 

mining,  36 

storage,  99 

strength  of,  51 
Column  pipe,  359 
Combustion  of  explosives,  426 

of  fuel,  88-96 
Compartments,  515 
Composition  of  explosives,  674 
Compound  cylinders,  108 
Compressed  air,  275,  408,  424 
efficiency  of,  335 
loss,  328 
transmission     of    power    by, 

334 

Comstock  mine,  522-567 
Condensers,  108-112 
Conductors,  175-177 
Conical  drum,  128 

hoist,  143 
Consumption  of  air,  96,  447 

of  fuel,  98 

of  powder,  677 

of  steam,  117,  355 

of  timber,  546 
Continuous  system  of  drilling,  509, 

654 

Cooke  fan,  422 
Cooling  air,  317 
Converters,  179 


INDEX. 


717 


Core,  508 

drill,  606 

lifter,  608 

Corliss  engine,  106,  114 
Cornish  pump,  348 
Counterbalance,  146 
Counterpoise,  Koepe,  146 
County  of  Durham  system,  63 
Cradle-dump,  248 
Creep,  390 
Cribbing,  525,  572 
Cross-cut,  22 
Cross-overs,  244 
Culm,  68 

Cummings'  system,  335 
Curb,  526 
Current,  alternating,  169 

capacity,  167 

continuous,  169 

direct,  169 
Curvature,  264 

radius  of,  201 

Cushier's  system  of  pumping,  355 
Cut-off,  104,  118 

automatic,  no 
Cyclone  prospector,  606 

DAMPS,  391,  394,  477,  493 

Darlington  drill,  394,  470,  643 

Davy  lamp,  471 

Dead- work,  24 

Deane  pump,  363 

Death-rate,  485 

De  Laval  wheel,  119 

Depression  produced  by  fans,  439 

Depth  of  holes,  626 

of  mine,  limiting,  504 

of  shafts,  505 
Derrick,  129,  210 
Designing  hoisters,  150 
Detaching-hook,  217 
Detonators,  673,  683 
Diamond  drill,  606 
Dick  lamp,  473 
Diesel  motor,  124,  370 
Diffusion  of  gas,  397 
Dimension-stone,  29,  623,  635 
Discipline  in  mines,  487 
Discovery,  1 5 
Ditches,  192 
Divining-rod,  8 
Doble  water-wheel,  190 
Dog-holes,  465 
Doors,  229,  462 
Double  entry,  19,  58 

hand-work,  627 
Draught,  99,  402,  410 
Drag  of  mine,  403 
Draw-bar  pull,  270 


Drifts,  26,  258,  571 
Drilling,  623 

by  Diamond  drill,  606 

by  hand-auger,  619 

by  power-drill,  637 

by  spring-pole,  597 
Drill-hole  deflections,  6n 
Drinker,  H.  W.,  583 
Drums,  130-142,  158 
Dualine,  680 
Dumb  channel,  402 
Dump,  248 
Durability,  rope,  203 
Dust  explosions,  494 
Duty,  365 
Dynamite,  679 
Dynamo,  171 

ECONOMIZERS,  92,  93 
Electric  coal-cutter,  664 

drill,  619,  655 

firing,  650,  691 

fuse,  692 

hoister,  162 

lamps,  481 

locomotive,  277 

motive  force,  167 

motor,  182,  381 

pump,  375 

signalling,  214 

symbols,  174 

transmission  of  power,  1 78 

units,  167 

wires,  164,  174 

Elevation  of  the  outer  rail,  264 
Endless  cable,  289,  298 

end  lines,  16 
End  on,  52 

Engine  capacity,  154,  158,  276,  280, 
282 

foundation,  117 

haulage,  290 

horse-power  of,  88,  115,  312,  373 

plane,  283 

plant,  87 

second  motion,  133 

speed,  114 

Engines,  classifications,  102,  108 
English  system  of  tunnelling,  583 

and  French  measures,  714 
Entry,  double,  19,  69 

ventilation,  405,  406 
Equivalent  orifice,  403,  430 
Exhaust  fan,  423 
Expansion-bits,  639 

-joints,  383 

of  air,  334 

of  steam,  102 
Exploitation,  34 


7i8 


INDEX. 


Exploratory  work,  10 
Explosions,  499 

force  of,  395 

barometric  relations,  396 
guidance  after,  475 
precautions  against,  474 
Explosive,  compressed  air,  424 

currents,  477 

definition  of,  672 

gas,  393 
Explosives,  671 

accidents  with,  67,  492 

flame  from,  682 

gases  from,  492,  68 1 

storage  of,  680 
Extinguishing  fires,  497 
Extra-lateral  rights,  13 
Eye,  see  Pick. 

FABRY  fan,  422 
Face  on,  52 
Face-entry,  59 
Fahrkunst,  253 
Fainting  in  shaft,  310 
Falls  of  ground,  489 

of  roof,  43,  489,  547 
Fan,  chimney,  100 

closed- running,  425 

design,  435 

efficiency,  434 

Guibal,  429 

relation  between  mine,  432 

theory,  429 

velocity,  427 

ventilation,  420 
Fault,  28 

Federal  mining  laws,  14 
Feed-water,  89 
Ferranti  electric  system,  97 
Fiery  mines,  447 
Filling  system.  65,  82 
Fire,  causes  of,  495 
Fire-damp  detection,  391,471,477, 

49.3 

Fire-setting,  616 
Firing,  barrel  system,  428 

electric,  650,  691 
Firing,  hand,  97 

-needle,  427 

First-motion  engine,  77,  430 
Fissure  veins,  5 
Flameless  explosive,  682 
Flash-point,  240 
Flat  rope,  198 

Fleuss  diving  apparatus,  501 
Float,  ii 
Flue-gases,  95 
Flushing.  68 
Foot- wall,  78 


Force-pump,  348 
Forcite,  432 
Forepoling,  533 
Forfeiture,  15 
Forge,  630 
Framing-machines,  574 

-tools,  572 
Free  air,  303 
Freezing  system,  543 
French  and  English  measures,  714 
Frequency,  173 

Friction,    coefficient    of,    193,   329, 
441 

-gear,  136 

-hoisters,  136 

of  air,  329 

of  cars,  238 

of  water,  189 
Fuel  consumption,  88-99 

-valve,  96 
Fulminate,  676 
Furnace  ventilation,  418 
Furnaces,  417 
Fuse,  electric,  676,  692 
Fusee,  76,  676 

GAD,  624 

Galleries,  dimensions  of,  26,  560 

Gangue,  4 

Gangway,  26,  258,  560 

centre  props,  555 
Gas  effects  upon  life  or  flame,  392 

occlusion,  391 
Gas-engine,  124 
Gaseous  mines,  61,  387 
Gases,  386 

explosive,  393 

from  explosions,  394,  493 

from  powder  combustion,  68 1 
Gauge  pressure,  308 

of  track,  262 
Geological  maps,  n 

theories,  5-8 

German  system  of  tunnelling,  584 
Giant  powder,  680 

nozzles,  31,  377 

duty  of,  32 
Glossary,  696 
Goaf,  401 ,  454 
Gob,  40,  66 

road,  44.  389,  557 
Governors,  109 
Grade  of  drift,  257 
Grate  area.  89 
Grip  drill,  618 
Guibal  fan.  423 
Guides.  222 
Gunboats,  224 
Guncotton,  673 


INDEX. 


719 


HAASE'S  system,  542 
Half-end  on,  52 
Hallidie  tramway,  300 
Hammer,  627 
Hardening  steel,  631 
Harrison  coal-cutter,  659 
Haulage,  257,  268,  276,  284 

-ways,  45-  457 
Heading,  259 
Head-gear,  209 
Heater,  93 

Hepplewite-Gray  lamp,  399,  473 
Hoisting  conveyances,  196 

-engine,  127-133 

-rope,  196 

speed,  512 

water,  338 
Holing,  612 
Hollenback  shaft,  523 
Hooks,  217 
Horse-power,  88,  115,  1 68,  321-328, 

382 

Housing  of  plant,  18,  90,  210,  431 
H -piece,  340 
Hudson  tramway,  301 
Hydraulic  engine,  372 

feed,  610 

mining,  30,  382 

ram,  338 

shield,  592 

wedge,  623 

Hydrogen,  sulphuretted,  289 
Hydrometer,  240 
Hyperian  logarithms,  314 

ILLUMINATION,  468,  469 
Inclined  planes,  284 
Inclines,  24,  69 
Indicated  horse-power,  115 
Indicators,  103,  214,  320,  361 
Induction  motor,  183 
Ingersoll  coal-cutters,  66 1 

compressors,  324 
Internal  gear,  136 
Intersecting  veins,  6,  16,  20 
Inysutse  wheels,  187 
Iron  in  tunnels,  592 

lining,  562 

-mines,  36,  72,  535 
Isothermal  compression,  310—312 

JARS,  598 
Jaws  of  rooms,  48 
Jeffrey  coal-cutter,  66 1 
Joints  of  timber,  514 
Juggler-way»,  64 
Jumper,  625 

KEEPS,  230 


Keystone  prospecting-machine,  606 

Kick  back,  244 

Kicking  down  a  hole,  599 

Kind-Chaudjon  process,  539 

Kirving,  see  Underholing. 

Knight  wheel,  188 

Koepe's  system  of  winding,  146 

Kutter's  formula,  192 

LADDERS,  249 

Lagging,  561 

Lamps,  394,  398,  399,  471,  477,  479 

Landings,  229,    265,    524,    573 

Latches,  236 

Laths,  533,  589 

Laws,  State  mining,  14,  210 

U.  S.  mining,  14 
Lemiele  fan,  423 
Levels,  distance  between,  25,   71, 

258,  560,  577 
Lewising,  637,  676 
Lift,  height  of,  25,  71 

-pumps,  345 
Limit  of  mining,  504 
Line  of  least  resistance,  686 
Lippman's  drill,  542 
Locked-wire  rope,  193 
Locomotives,  277 
Lode,  definition  of,  5,  9,  13,  18 
Logarithms,  hyperbolic,  314 
Long-hole  process,  509,  654 
Long  horn,  52,  47 
Long  tunnels,  581 
Longwall,  664 
Lubricating-oil,  238 

MACHINE  vs.  hand-work,  509,  648, 

659 

Magic  wand,  8-13 
Magneto  machine,  692 
Man-engine,  252 
Manganese  ores,  249 
Man-hole,  see  Mill-hole,  249 
Manometric  depression,  425 
Manual  haulage,  267 
Mapping,  n,  498 
Marsault  lamp,  472 
Marsh-gas,  390 
Masonry  in  mines,  529,  563 
Mather  and   Platt  boring  system, 

60 1 

Mean  effective  pressure,  105,  118 
Mechanical  ventilators,  420 
Mercury  veins,  4 
Meyer  valve,  in,  322 
Mica,  4 

Mil,  circular,  169 
Mill-hole,  36,  74,  551,  574 


720 


INDEX. 


Mine  buildings,  18,  90,  210,  431 
Mine,  definition  of,  3 
Miner's  inches,  382,  714 

tools,  616 
Mining,  economy  of,  46,  54 

in  soft  ore,  75 

in  thick  seams,  61 

.n  thick  veins,  75 

in  thin  seams,  38 

in  thin  veins,  38 

laws,  14 

limit,  504 

terms,  696 

claim,  14-16 

retreating,  38 
Moil,  624 

Moment  of  an  engine,  155 
Moss-box,  541,  604 
Motive  column,  413,  441 
Motor,  electric,  182 
Mueseler  lamp,  472 
Multiple  circuits,  457 

NARROW  work,  57 
Native  metal,  2 

ventilation,  409 
Nickel  ores.  4 
Nitrogen,  388 
Nitroglycerine,  678 

storage  of,  427 
Norwalk  air-compressor,  325 
Nottingham  system,  47 

OCCLUSION  of  gas,  391 
Ohm,  1 68 
Oil-engine,  124 
Oil,  illuminating-,  469 

lamp,  468 

lubricating-,  238 

prospecting  for,  601 
Oil-well  rig,  598 
Oilers,  self-,  238 
Open  cut,  34 

-running  fans,  424 
Ore  deposition,  4 

bucket,  218,  308 
definition  of,  3 

-shoot,  7 

Outbursts  of  gas,  391 
Outcrop  of  coal,  10 

of  veins,  10,  14,  18 
Outlet,  single,  507 
Overcast,  460 
Overhand  method,  70—73 
Overloaded  engine,  62 
Overwinding,  216-492 
Oxygen,  388 

PACK-WALL,  40 


Panel  system,  Brown's,  61,  447 

Parson's  steam-turbine,  119 

Patenting  a  claim,  16 

Pay  streak,  6-25 

Peat-mining,  31 

Pelton  wheel,  190 

Pentice,  510 

Petroleum,  monograph  on,  12 

Phosphate  rock,  4 

Pick,  621 

Pickets,  spilling,  390 

Pieler  gas-detector,  473 

Pike,  see  Pick, 

Pillar  and  room,  38-48 

Pillars,  robbing  of,  57,  65,  392 

sizes  of,  27-53,  73,  447,  531 

waste  in  recovering,  42 
Pinch,  6-26 

Pipe  sizes,  190,333,342,359 
Pipes,  iron,  194,  333,  359 
Piston  speed,  367 

pump,  354 
Planimeter,  106 
Pneumatic  system,  537 
Poetsch  sinking  system,  543,  577 
Pole-pick,  621 
Poling,  588 
Post  drill,  618 
Powder,  495,  671 

accidents  with,  681 

consumption  of,  677 
Power  plants,  87 

drills,  641 

transmission,  101,  178,  296,  335 

-driven  pumps,  374 
Preparatory  work,  18 
Pressure  of  air,  308 
Prop,  491,  548, 668 
Prospecting    by    boring,    10,  504, 

597 

by  wand,  813 

by  witchery,  13 

in  massive  rock,  10 

in  stratified  rock,  n 

surface,  10 
Pumps,  336 

centrifugal,  377 
Pump-bob,  352 
Pump,  Bull,  347 

Cook,  345 

Cornish,  348 

duty  of,  365 

electric,  375 

hydraulic,  363 

rod,  348 

rotary,  379 

sinking,  356 

-valve,  343,  355 
Pulsometer,  383 


INDEX. 


721 


Punch-drills,  597 

8UARRY,   29,   35,  635 
uarrying     dimension-stone,      29, 
623-636 

RAILS,  261 

Ram,  636 

Ramsey  caging,  244 

Rand  compressors,  323 

drills,  639 
Rateau  fan,  425 
Recorder,  428 
Reels,  144-158 
Regulator-doors,  463 
Reheating  air,  334 
Reserves,  26 

Resistance,  402,  419,  445,  448 
Retreating,  mining,  44 
Reversible  engine,  112 
Riedler  pump,  375 
Rifled  holes,  645 
Rheostat,  184,  279 
Rivet,  pipe-,  340 
Robbing  pillars,  57,  65,  392 
Rock-chute  mining,  63 
Rock  pressure,  49 

-shaft,  82 
Room,  54 
Root-blower,  422 
Rope,  152 

-drilling,  196 

flat,  284 

haulage,  284 

preservation,  203 

round,  201 

strength  of,  153,  199,  297 
Rotary  bar-cutter,  660 

effort,  118 

converters,  177 

pumps,  397 
Rupturing  effect  of  explosives,  673 

SAFETY  appliances,  265 

-boiler,  91 

-cage,  226 

-catch,  218 

-doors,  464 

explosives,  682 

-lamp,  471 
Sag  of  rope,  298 
Salt-mining,  31 
Sand-pump,  602 
Scale  in  boiler,  92 
Schiele  fan,  230 
Schmidt's  rule,  28 
Scotch  system,  52 
Schram's  drill,  640 
Second-motion  engines,  33 


Self-acting  plane,  55,  282 
Self-dumping  cages,  231 
Self-recorder  for  fans,  428 
Selvage,  5 

Sergeant  drill,  641,  660 
Shaft,  504 

auxiliary,  580 

-bottoms,  531 

compartment,  507-515 

-pillars,  27,  42,  73,  531 

rectangular,  507 

round,  530 

shape  of,  530 

sinking,  22,  513,  543 

site  of,  20,  69,  505 

size  of,  23,  512 

-timbering,  512 
Shaw's  gas-detector,  399 
Shearing  coal,  663 
Sheave,  210,  290,  296 
Shode,  ii    . 
Shoot  ore,  7 
Shorthorn,  52 
Shovels,  621 
Shutter  on  fans,  427 
Signals,  213 
Sills,  574 

Simultaneous  firing,  650,  691 
Single  chute,  64 

entry,  504 

-handwork,  627 
Sinking,    continuous   process,    509,. 

.     654     . 

in  running  ground,  534 

Kind-Chaudron  process,  537 

Poetsch's  system,  543 

pump,  356 
Siphon,  384 
Skips,  223 
Slitter,  see  Pick. 
Slope,  579 

-cage,  228 

-carriage,  227 

openings,  21 

railway,  284 
Slow  powder,  673 
Sludger,  539,  605 
Smokeless  powder,  684 
Snap-hooks,  206 
Soft  ground,  drifting  in,  611 

timbering  in,  566 
Spades,  406 
Speed  of  drilling,   610,   626,  648- 

of  haulage,  268,  287,  292 

of  hoisting,  151 

of  pumping,  364-367 

of  ventilating  current,  443,  455; 
Spilling,  see  Poling. 
Splicing  rope,  108 


722 


IXDEX. 


Splitting  air-current,  447 

Spontaneous  combustion,  497 

Spoon, 627 

Sprag,  43,  258 

Spring-hole  drilling,  2,  599 

Spudding,  599 

Square  sets,  75,  518,  568 

work,  67 
Squeeze,  390 
Squib,  617,  672, 692 
Statutory  provisions,  486 
Stables,  underground,  269 
Stalls,  40 
Steam  as  extinguisher,  102 

-boiler,  88 

coal;  97 

condenser,  108 

expansion,  105 

-jet  ventilation,  100 

pressure,  117,  102 

-pump,  361 

-shovel  30,  35 

-turbine,  119 
Steel,  definition,  631 
Stent,  627 

Stephenson  lamp,  472 
Stockwerke,  4-6 
Stokers,  98 
Stone  shaft,  83 
Stope,  39,  70,  74,459 

height  of,  25,  71 
Storage  of  powder,  680 

battery,  166-281 
Stowing,  see  Filling. 
Stripping,  36 
Stull,  74,  548 
Stump-pillars,  47,  57 
Sty  the,  see  Carbonic  acid. 
Sub-drifting,  80 
Suction  height,  367 
Suction-pipe,  344 
Sullivan  machine,  663 
Sulphuretted  hydrogen,  389 
Surface  examination,  12 

TAIL  out,  ii 

Tail-rope  haulage,  148,  285 

hoisting,  147 
Tamping,  675 
Taper- rope,  199 
Tappets,  639 
Telephones,  215 
Temper- screw,  60 1 
Temperature  of  deep  mines,  408, 

5°5 

Tempering  steel,  632 
Tension-wheel,  292,  296 
Testing  for  gas,  398 
Three- wire  system,  178 


Throttling  governors,  109 
Throughs,  465 
Timber-joints,  516,  542 

preservatives,  546 
Timbering  gangways,  555 

rooms,  43 

shafts,  514 

slopes,  28 

stopes,  75 
Timbering,  iron  for,  562 

methods  of,  77,  514,  554 

principles  of,  547 
Tin  ores,  4 

Tipple,  18,  88,  209,  242 
Tools,  594-615 
Top-slicing.  79 
Traction,  animal,  68 

locomotive,  270 
Tractive  force,  271 
Tramway,  267,  299 
Transmission,  329 
Traverses,  4,  31 
Trepan,  538 
Triger's  method,  537 
Trompe,  421 
Tubes,  535 

Tubing,  528,  535-540 
Tubing  wells,  603 
Tunnel,  584 

long,  605 

openings  by,  22 

progress,  606 
Tunnelling  methods,  607 
Turbines,  119,  189 
Two- wire  system,  178 

ULTIMATE  source  of  mineral,  8 
Undercast,  461 
Underhand  stope,  70 
Underholing,  42,  550,  620,  658 
Upcast,  509 
Upraise,  see  Mill-hole. 
U.  S.  mining  laws,  14 

VALVES  of  air-compressors,  326 

of  drills,  638 

of  engines,  106 

of  pumps,  342-355 
Vein,  3 

definition  of,  5 

formation,  7 
Ventilating  current,  455 
Ventilation,  386,  405,  406 

amount  of  air,  400 

natural,  409 

of  breasts,  286 

of  tunnels,  161,  323,  380 

power  required,  450 


INDEX 


Viscosity  of  oils,  240 
Voltage,  167,  18 1 

WAC.OX-BOXES,  236 

-breast,  49 
Wall-plates,  514 
Water-bailers,  335 

-gauge,  401,  415,  449 

heaters,  93 

-level  theory,  8 

-pressure  engines,  371 

-purifiers,  94 

-wheels,  186 
Watt,  1 68 
Weights  and  measures,  714 

of  various  substances,  411,   714 


Welding,  630 
Westphalian  system,  66 
Wheels  of  cars,  238 
Whim,  128 
White  damp,  389 
Whiting  system,  148 
Winding,  see  Hoisting. 
Windlass,  127 
Wings,  230 
Winze,  70 
Wire  rope,  197,  296 
Wires,  connecting,  164,  175,  691 
Wolf  lamp,  474 
Working  barrel,  344 
Work  of  compressing  air,  314,  321, 
334 


UNIVERSITY  OF  CALIFORNIA  LIBRARY 

Los  Angeles 
This  book  is  DUE  on  the  last  date  stamped  below. 


OCT  2  9  1952 

MAY  1 2  1958 

a  a  o  Ktcn 

DEC     5  v 
ncr.    5  RECO 


I  2  R£C*0 


Form  L9-25m-8,'46  ( 9852 ) 444 


A     000226556 


TN 
145 
I25m 
1905 

peering 
Library 


STACK 

SEP     73 


